CA1106620A - Method for recovering zinc - Google Patents
Method for recovering zincInfo
- Publication number
- CA1106620A CA1106620A CA283,129A CA283129A CA1106620A CA 1106620 A CA1106620 A CA 1106620A CA 283129 A CA283129 A CA 283129A CA 1106620 A CA1106620 A CA 1106620A
- Authority
- CA
- Canada
- Prior art keywords
- leaching
- zinc
- residue
- mixed oxides
- neutral
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Expired
Links
Landscapes
- Manufacture And Refinement Of Metals (AREA)
Abstract
ABSTRACT OF THE DISCLOSURE
A method for recovering zinc from oxidic zinc ores or from calcined-zinc sulphide ore concentrates, or both, the ores or concentrates being leached, and leaching residues thus obtained being subjected to a pyrometallurgical volatilization process. The method comprises the steps of: neutral leaching calcined zinc material so as to produce a leaching residue having a high content of undissolved zinc; drying the leaching residue by bringing the residue into contact with hot gases under reduc-ing conditions and in a state of suspension; volatilizing and precipitating the zinc and remaining volatile metal compounds in the form of metal oxides; and finally subjecting the mixed oxides to neutral leaching.
A method for recovering zinc from oxidic zinc ores or from calcined-zinc sulphide ore concentrates, or both, the ores or concentrates being leached, and leaching residues thus obtained being subjected to a pyrometallurgical volatilization process. The method comprises the steps of: neutral leaching calcined zinc material so as to produce a leaching residue having a high content of undissolved zinc; drying the leaching residue by bringing the residue into contact with hot gases under reduc-ing conditions and in a state of suspension; volatilizing and precipitating the zinc and remaining volatile metal compounds in the form of metal oxides; and finally subjecting the mixed oxides to neutral leaching.
Description
The invention relates to a method for recovering zinc from oxydic zinc ores and/or from calcined zinc-sulphide-ore con-centrates, the said ores or concentrates being leached and the resulting leaching residues being sub]ected to a pyrometallur-gical volatilization process.
Industrial recovery of zinc by leaching and electrolysis has acquired increasing significance since its discovery at the beginning of this century, especially after it became known that the ore concentrates best suited to the purpose were those con-taining large amounts of zinc and very few contaminants. In therecovery of zinc by electrolysis, "contaminants" are mainly iron, silicic acid, arsenic, antimony, cobalt, nickel, germanium, chlor-ine and fluorine. However, since these elements, almost without exception, are found with zinc blende, the main zinc ore, it was found necessary, in the past 50 years, to master the industrial application of zinc electrolysis, in order to render these "con-taminants" harmless, since they caused trouble both in leaching and in electrolytic separation.
Thus, the economics of zinc recovery by electrolysis, is largely dependent upon eliminating the said contaminants as far as possible during the calcining and leaching processes.
At the start of the industrial application of the zinc-electrolysis process, when high-grade and pure zinc ores and ore concentrates were readily available, and when thoughtless pollution by the discarding of residues involved no substantial economic penalties, people were content with conventional neutral leach-ing. However, with the increasing need for treating ores con-taining large amounts of iron, and with increasing fluctuation in the composition of the ores, it became difficult, and even to some extent uneconomical, to use the electrolytic process.
This sequence of problems begins with the primary leach-ing of the calcined material used, since when the oxydic zinc carrier is treated with dilute sulphuric acid, the iron which previously passed into solution is precipitated as ferric hy-droxide at between pH 3 and 4, as the acidity decreases. At the same time, the iron hydroxide, as it flocculates, precipi-tates a large proportion of the contaminants, such as Asj Sb and Ge. ~Iowever, in order -to achieve this decisively important effect, it is necessary to add an excess of calcined material dependent upon the volume of contaminants, the inevitable re-sult of which is that the zinc oxide therein contained is no longer dissolved. This produces yields of between 70 and, at most, 90% of zinc, the leaching residues containing unacceptably large amounts of undissolved zinc.
In order, therefore, to eliminate the expense of metal losses, it was proposed to improve the wet metallurgical process to produce yields in excess of 90% in the leaching plant. In the past, this was achieved by introducing a plurality of con-secutive leaching stages with modified leaching and precipitating conditions, with some very high acid concentrations and tempera-tures, and with the addition of gaseous oxidizing agents. This method became possible especially after it was realized that large amounts of co-dissolved iron could be precipitated from the solution, under certain conditions, in the form of jarosite or goethite in an easily filterable residue.
~ owever, the disadvantage of this method is that the mul-ti-stage leaching operation involves large investment and operat-ing costs. For example, the leaching fluid, the concentration of which for the electrolytic process is predetermined by theoreti-cal principles, must be raised to the necessary much higher con-centration in a separate evaporating unit. Furthermore, since the leaching operation is carried out in multiple layers, the leaching-plant requires much more complex equipment, and more sophisticated supervision and control.
Finally, one particularly serious disadvantaye is that the waste from the jarosite and/or goethite process causes con-siderable pollution problems, since the harmful substances con-tained therein, for example, compounds of arsenic, antimony, thallium, bismuth, etc., are easily washed out by rain water, and thu~s pollute the subsoil water if they are dumped in the open.
In these days, therefore, these multi-leaching-stage installations are burdened with such costs for disposing of the residues that the former economic advantages of the purely wet-metallurgy treat-ment process are no longer applicable.
Pyrometallurgical treatment of the leaching residues represents a totally different approach, and a whole series of processes has been proposed, tested, and even used on a large in-dustrial scale. These processes range between rolling and bloom-rolling furnaces and a very wide variety of shaft furnaces, semi-shaft furnaces, and sintering furnaces, with or without super-charging. The objective was always to improve the zinc yield from the hydrometallurgical leaching process by combining it with the pyrometallurgical volatilization process, the leaching and vol-atilization being united in a cyclic process.
The known principle upon which this process is based is as follows: after the leaching residue has been dried, carbon is added, and the whole is mixed with slag-forming additives. This mixture is melted down under reducing conditions at temperatures ; above 1000C. which causes the zinc content to volatilize as a metal vapour. This i9 oxidized in a flow of gas, and is finally recovered in dust-collecting units. The non-volatile and slag-forming components produce a molten phase of rock and slag. The high-zinc-content, oxide sulphate dust recovered is returned to the leaching stage.
At the beginning of the 50's, this process was regarded with disfavour because the dry leaching residue is in the form of a very fine powder and requires, for proper heat-treatment, a spe-cial pre-treatment such as pelletizing, briquetting, or sintering, which is highly expensive and complex. Furthermore, all of the above-mentioned furnace installa-tions have restricted outputs and are tailored to former zinc-smelting installations with capa-cities of between 50 000 and 80,000 tons per annum. Some of the heat treatments also require expensive sources of energy, such as metallurgical coke or oil. Since this was also a multi-stage ther-mal process for the recovery of metal, in the various stages of which it was impossible to prevent the formation of dust, and since dilute waste gases did not support industrial processing, these variants of the process revealed a series of disadvantages as production volume increased, thus preventing their economical application. They also cause pollution, have high operating and maintenance costs, and require considerable capital investment.
Based on the foregoing knowledge, it is the purpose of the present invention to simplify once more the leaching pro-cess and to link it rationally with a heat-treatment process for the leaching residue such that the resulting waste can be discard-ed with absolute safety, while the overall yield of zinc is opti-mal. A simple single-stage leaching process is used, and this converts the zinc, which is easily dissolved out of the calcined material, into the fluid phase at minimal cost. The resulting zinc-rich solution residues, with all of their secondary metals, are then subjected to a thermaL process having none of the disad-vantages mentioned above.
As already indicated, one characterlstic of the dried leaching residue is its eY~treme fineness. The invention assumes that this characteristic can be utilized economically for ad-vantageous processing. The purpose of the invention is achievedin that neutral leaching of the calcined zinc material produces a leaching residue having a high content of undissolved zinc~
Industrial recovery of zinc by leaching and electrolysis has acquired increasing significance since its discovery at the beginning of this century, especially after it became known that the ore concentrates best suited to the purpose were those con-taining large amounts of zinc and very few contaminants. In therecovery of zinc by electrolysis, "contaminants" are mainly iron, silicic acid, arsenic, antimony, cobalt, nickel, germanium, chlor-ine and fluorine. However, since these elements, almost without exception, are found with zinc blende, the main zinc ore, it was found necessary, in the past 50 years, to master the industrial application of zinc electrolysis, in order to render these "con-taminants" harmless, since they caused trouble both in leaching and in electrolytic separation.
Thus, the economics of zinc recovery by electrolysis, is largely dependent upon eliminating the said contaminants as far as possible during the calcining and leaching processes.
At the start of the industrial application of the zinc-electrolysis process, when high-grade and pure zinc ores and ore concentrates were readily available, and when thoughtless pollution by the discarding of residues involved no substantial economic penalties, people were content with conventional neutral leach-ing. However, with the increasing need for treating ores con-taining large amounts of iron, and with increasing fluctuation in the composition of the ores, it became difficult, and even to some extent uneconomical, to use the electrolytic process.
This sequence of problems begins with the primary leach-ing of the calcined material used, since when the oxydic zinc carrier is treated with dilute sulphuric acid, the iron which previously passed into solution is precipitated as ferric hy-droxide at between pH 3 and 4, as the acidity decreases. At the same time, the iron hydroxide, as it flocculates, precipi-tates a large proportion of the contaminants, such as Asj Sb and Ge. ~Iowever, in order -to achieve this decisively important effect, it is necessary to add an excess of calcined material dependent upon the volume of contaminants, the inevitable re-sult of which is that the zinc oxide therein contained is no longer dissolved. This produces yields of between 70 and, at most, 90% of zinc, the leaching residues containing unacceptably large amounts of undissolved zinc.
In order, therefore, to eliminate the expense of metal losses, it was proposed to improve the wet metallurgical process to produce yields in excess of 90% in the leaching plant. In the past, this was achieved by introducing a plurality of con-secutive leaching stages with modified leaching and precipitating conditions, with some very high acid concentrations and tempera-tures, and with the addition of gaseous oxidizing agents. This method became possible especially after it was realized that large amounts of co-dissolved iron could be precipitated from the solution, under certain conditions, in the form of jarosite or goethite in an easily filterable residue.
~ owever, the disadvantage of this method is that the mul-ti-stage leaching operation involves large investment and operat-ing costs. For example, the leaching fluid, the concentration of which for the electrolytic process is predetermined by theoreti-cal principles, must be raised to the necessary much higher con-centration in a separate evaporating unit. Furthermore, since the leaching operation is carried out in multiple layers, the leaching-plant requires much more complex equipment, and more sophisticated supervision and control.
Finally, one particularly serious disadvantaye is that the waste from the jarosite and/or goethite process causes con-siderable pollution problems, since the harmful substances con-tained therein, for example, compounds of arsenic, antimony, thallium, bismuth, etc., are easily washed out by rain water, and thu~s pollute the subsoil water if they are dumped in the open.
In these days, therefore, these multi-leaching-stage installations are burdened with such costs for disposing of the residues that the former economic advantages of the purely wet-metallurgy treat-ment process are no longer applicable.
Pyrometallurgical treatment of the leaching residues represents a totally different approach, and a whole series of processes has been proposed, tested, and even used on a large in-dustrial scale. These processes range between rolling and bloom-rolling furnaces and a very wide variety of shaft furnaces, semi-shaft furnaces, and sintering furnaces, with or without super-charging. The objective was always to improve the zinc yield from the hydrometallurgical leaching process by combining it with the pyrometallurgical volatilization process, the leaching and vol-atilization being united in a cyclic process.
The known principle upon which this process is based is as follows: after the leaching residue has been dried, carbon is added, and the whole is mixed with slag-forming additives. This mixture is melted down under reducing conditions at temperatures ; above 1000C. which causes the zinc content to volatilize as a metal vapour. This i9 oxidized in a flow of gas, and is finally recovered in dust-collecting units. The non-volatile and slag-forming components produce a molten phase of rock and slag. The high-zinc-content, oxide sulphate dust recovered is returned to the leaching stage.
At the beginning of the 50's, this process was regarded with disfavour because the dry leaching residue is in the form of a very fine powder and requires, for proper heat-treatment, a spe-cial pre-treatment such as pelletizing, briquetting, or sintering, which is highly expensive and complex. Furthermore, all of the above-mentioned furnace installa-tions have restricted outputs and are tailored to former zinc-smelting installations with capa-cities of between 50 000 and 80,000 tons per annum. Some of the heat treatments also require expensive sources of energy, such as metallurgical coke or oil. Since this was also a multi-stage ther-mal process for the recovery of metal, in the various stages of which it was impossible to prevent the formation of dust, and since dilute waste gases did not support industrial processing, these variants of the process revealed a series of disadvantages as production volume increased, thus preventing their economical application. They also cause pollution, have high operating and maintenance costs, and require considerable capital investment.
Based on the foregoing knowledge, it is the purpose of the present invention to simplify once more the leaching pro-cess and to link it rationally with a heat-treatment process for the leaching residue such that the resulting waste can be discard-ed with absolute safety, while the overall yield of zinc is opti-mal. A simple single-stage leaching process is used, and this converts the zinc, which is easily dissolved out of the calcined material, into the fluid phase at minimal cost. The resulting zinc-rich solution residues, with all of their secondary metals, are then subjected to a thermaL process having none of the disad-vantages mentioned above.
As already indicated, one characterlstic of the dried leaching residue is its eY~treme fineness. The invention assumes that this characteristic can be utilized economically for ad-vantageous processing. The purpose of the invention is achievedin that neutral leaching of the calcined zinc material produces a leaching residue having a high content of undissolved zinc~
2~
This leaching residue is then dried, is brought into contact l"ith hot gases under reducing conditions and in a state of suspension.
The zinc and the remaining volatile metal compounds are volatil-ized and precipitated in the form of mixed oxides, whereas the non-volatile metals and metal compounds occur in the form of slag or rock. Finally, the mixed oxides are subjected to neutral leaching.
More specifically, according to the invention, there is provided a method for recovering zinc from oxidic zinc ores or from calcined zinc-sulphide ore concentrates, or both, the ores or concentrates being leached and leaching residues thus obtain-ed being subjected to a pyrometallurgical volatilization process, comprising the steps of neutral leaching calcined zinc material with sulphuric acid solution so as to produce a leaching residue having a high content of undissolved zinc' drying the leaching residue by bringing the residue into contact with hot gases under reducing conditions and in a state of suspension; volatil-izing said residue at a temperature of about 1~50C and precip-itating the zinc and remaining volatile metal comp~unds in the form of mixed oxides, whereas non-volatile metals and metal com-pounds occur as slag or rock, and, finally, subjecting the mixed oxides to neutral leaching with a wea~ sulphuric acid solution.
According to one preferred embodiment of the method relating to the present invention, the neutral leaching of the calcined zinc-ore material is carried out in a single stage. In this connection, it is also desirable to carry out the neutral leaching until a degree of acidity corresponding to a p~ value of . about 5, preferably 5.5, is reached,and until the iron, which has passed into solution, precipitates, and until the metal ions of other accompanying metals, which have passed into solution, are al.so precipitated with the flocculating iron compounds.
2~
A preferred embodiment of the process according to the invention is characterized by pyrometalLurgical treatment of the leaching residue in a fusion cyclone.
The invention thus combines neutral leaching with sim-ple equipment and a simple process. As well, the invention also overcomes a widespread expert prejudice with a pyrometallurgical process, regarded as both functionally and economically optimal, consisting of a "solid-in-gas" reaction in which the solid is in the form of a very fine powder, thus eliminating the disadvantages associated with each of these two processes per se.
The pyrometallurgical process according to the invention makes it possible to heat the material, which is so fine that it can trickle, to a smelting temperature of about 14S0C, so that the thermo-chemical process takes place with the solid/gas in the "freely floating" condition, over a very short period of time, with the best results. This illustrates the significant difference - 5a 6~
between the claimed reaction vessels and a rolliny furnace, for instance, which requires a 2 m3 furnace chamber for each ton/day whereas in a fusion cyclone, for example, a 1 m3 reaction chamber is enough for a daily throughput of 25 t. The reaction chamber ratio is thus about 50:1. This signifies a considerable saving in investment capital and space, in addition to the extremely intense, and therefor rapid reaction in the cyclone.
According to a further embodiment of the me-thod accord-ing to the invention, the mixed oxides obtained by the pyro-metallurgical method are subjected, separately from the calcinedzinc material, to neutral leaching in a weak sulphuric-acid solu-tion. This causes the main part of the zinc oxide, contained in a particu~rly easily-soluble :Eorm in the mixed oxides, to pass into solution, while the resulting leaching residue is enriched by the remaining valuable metallic substances.
Another advantage provided by the invention, is that with separate leaching and precipitation of the mixed oxides, this leaching and precipitation proceeds according to the leaching and precipitation capabilities of the different substances of value, in such a manner that thereis consecutive enrichment of those substances of value that are difficult to dissolve. In this connection, it is to the advantage of the process according to the invention that the mixed oxides be in the form of a collection of the finest partlcles with an optimum of active surface~ The mixed oxides therefore contain individual components in a particu-larly easily soluble form, a property which assists the leaching and precipitation.
Finally, this also has the advantage of directing the process according to the in~ention in such a manner that the mixed oxides are used, during primary leaching of the calcined zinc material, towards the end of the neutralizing, at a pH value of 2, preferably greater than 2. The invention is explained herein-after by means of an example.
Example Calcined zinc material, containing the following amountsof valuable metals, is placed in a leachiny vat equipped with an agita-tor:
Zn =47.52% S = 29.28%
Cd =0.20% Mn = 0.23%
Pb =1.08% As = 0.03%
Cu =0.44% Ag = 151 g/t Fe =11.80% Remainder = About 3.50%
(The"remainder" contains ballast substances in the form of SiO2, A12O3, CaO, MgO, etc.) The calcined material is ground to 70/O ~ 75~ (200 mesh) and is placed for leaching in a vat containing an agitator.
Neutral leaching starts with an initial content of about 115 g/l of free sulphuric acid, and is continued to a neutral point of p~I 5.5 for the zinc-sulphate solution. Calcined material is added to the acid solution until this pH value is reached. The leaching temperature lies between 50 and 70~C.
Concentrating the sludge water produces a residue of about 490 kg per ton of calcined material, i.e., 49%. The composi-tion of this residue (in relation to dry substance) is as follows:
ZnO = 14.8% CuO =0.71%
ZnS = 3.4% CdO =0.09%
ZnSO4 = 6.2% AS23 ~0.24%
ZnO.Fe2O3 = 50.6% n2 0.38%
Fe ~OH)3 = 8.7% g2 600 g/t PbSO4 = 5.7% Remainder = about 3.20%
(The "remainder" consists of SiO2, A1203, CaSO4 MgO, etc.).
This residue is dried to a fine powder. It is mixed with 30% by weiyht of finely-yround coke breeze (about 200 mesh~
and is fed uniformly, suspended in preheated air, into a fusion cyclone.
In addition to the coke breeze mentioned above, SiO2 and FeS2 are mixed with the finely-granular leaching residue as additives, for the purpose of forming slag and rock.
Upon leaving the fusion reactor, the gases and molten products are separated. The hot gases pass to an after-burning chamber where, with the addition of air, carbon monoxide and the volatile metal vapours are reoxidized, i.e., burned. These gases pass through a cooler into a filter from which the metal-oxide dusts, which are carried along and contain large amounts of the volatile components, are recovered in the form of mixed oxides.
The molten products are collected in a settling hearth where a slag/rock separation is carried out. The rock contains non-volatile metals, e.g., copper, nickel and noble metals in the enriched state, whereas the gangue, and most of the iron, pass into the slag.
rrhe mixed oxides are the products of the following chem-ical reactions:
2 _ ~ FeS + 1/2 S2 ~-C ~ 2 ~ C2 C2 + C _ _ ~ 2co ZnO + CO ~ Zn + CO2 ZnO Fe2O3 + 2CO ~ Zn + 2 FeO ~ 2 CO2 ZnSO4 + 4CO > ZnS + 4 CO2 ZnS + FeO + CO _ _ ) Zn + FeS + CO2 PbSO4 -~ 4 CO_ > PbS + 4 CO2 CdO + CO ~ Cd + C02 After-burning of the metal vapours produces:
Zn + 1/2 2 - ~ > ZnO
Cd + 1/2 2 ~ CdO
PbS ~ 3/2 2 ~ PbO + S02 PbS + 1/2 2 > PbS -~ 2 2 ) PbS04 The composition of the mixed oxide appearing in the dust collector is as follows:
ZnO = 75.3% 5dO = 0.9%
PbO = 13.3% 2 3 = 0.8%
The remainder, 9.7% consists of impurities carried 1.0 along,for example Fe203, C, SiO2 etc.
As previously described, the mixed oxides are sub-jected to leaching and precipitation, the constituents being relatively easily separated from each other.
.
This leaching residue is then dried, is brought into contact l"ith hot gases under reducing conditions and in a state of suspension.
The zinc and the remaining volatile metal compounds are volatil-ized and precipitated in the form of mixed oxides, whereas the non-volatile metals and metal compounds occur in the form of slag or rock. Finally, the mixed oxides are subjected to neutral leaching.
More specifically, according to the invention, there is provided a method for recovering zinc from oxidic zinc ores or from calcined zinc-sulphide ore concentrates, or both, the ores or concentrates being leached and leaching residues thus obtain-ed being subjected to a pyrometallurgical volatilization process, comprising the steps of neutral leaching calcined zinc material with sulphuric acid solution so as to produce a leaching residue having a high content of undissolved zinc' drying the leaching residue by bringing the residue into contact with hot gases under reducing conditions and in a state of suspension; volatil-izing said residue at a temperature of about 1~50C and precip-itating the zinc and remaining volatile metal comp~unds in the form of mixed oxides, whereas non-volatile metals and metal com-pounds occur as slag or rock, and, finally, subjecting the mixed oxides to neutral leaching with a wea~ sulphuric acid solution.
According to one preferred embodiment of the method relating to the present invention, the neutral leaching of the calcined zinc-ore material is carried out in a single stage. In this connection, it is also desirable to carry out the neutral leaching until a degree of acidity corresponding to a p~ value of . about 5, preferably 5.5, is reached,and until the iron, which has passed into solution, precipitates, and until the metal ions of other accompanying metals, which have passed into solution, are al.so precipitated with the flocculating iron compounds.
2~
A preferred embodiment of the process according to the invention is characterized by pyrometalLurgical treatment of the leaching residue in a fusion cyclone.
The invention thus combines neutral leaching with sim-ple equipment and a simple process. As well, the invention also overcomes a widespread expert prejudice with a pyrometallurgical process, regarded as both functionally and economically optimal, consisting of a "solid-in-gas" reaction in which the solid is in the form of a very fine powder, thus eliminating the disadvantages associated with each of these two processes per se.
The pyrometallurgical process according to the invention makes it possible to heat the material, which is so fine that it can trickle, to a smelting temperature of about 14S0C, so that the thermo-chemical process takes place with the solid/gas in the "freely floating" condition, over a very short period of time, with the best results. This illustrates the significant difference - 5a 6~
between the claimed reaction vessels and a rolliny furnace, for instance, which requires a 2 m3 furnace chamber for each ton/day whereas in a fusion cyclone, for example, a 1 m3 reaction chamber is enough for a daily throughput of 25 t. The reaction chamber ratio is thus about 50:1. This signifies a considerable saving in investment capital and space, in addition to the extremely intense, and therefor rapid reaction in the cyclone.
According to a further embodiment of the me-thod accord-ing to the invention, the mixed oxides obtained by the pyro-metallurgical method are subjected, separately from the calcinedzinc material, to neutral leaching in a weak sulphuric-acid solu-tion. This causes the main part of the zinc oxide, contained in a particu~rly easily-soluble :Eorm in the mixed oxides, to pass into solution, while the resulting leaching residue is enriched by the remaining valuable metallic substances.
Another advantage provided by the invention, is that with separate leaching and precipitation of the mixed oxides, this leaching and precipitation proceeds according to the leaching and precipitation capabilities of the different substances of value, in such a manner that thereis consecutive enrichment of those substances of value that are difficult to dissolve. In this connection, it is to the advantage of the process according to the invention that the mixed oxides be in the form of a collection of the finest partlcles with an optimum of active surface~ The mixed oxides therefore contain individual components in a particu-larly easily soluble form, a property which assists the leaching and precipitation.
Finally, this also has the advantage of directing the process according to the in~ention in such a manner that the mixed oxides are used, during primary leaching of the calcined zinc material, towards the end of the neutralizing, at a pH value of 2, preferably greater than 2. The invention is explained herein-after by means of an example.
Example Calcined zinc material, containing the following amountsof valuable metals, is placed in a leachiny vat equipped with an agita-tor:
Zn =47.52% S = 29.28%
Cd =0.20% Mn = 0.23%
Pb =1.08% As = 0.03%
Cu =0.44% Ag = 151 g/t Fe =11.80% Remainder = About 3.50%
(The"remainder" contains ballast substances in the form of SiO2, A12O3, CaO, MgO, etc.) The calcined material is ground to 70/O ~ 75~ (200 mesh) and is placed for leaching in a vat containing an agitator.
Neutral leaching starts with an initial content of about 115 g/l of free sulphuric acid, and is continued to a neutral point of p~I 5.5 for the zinc-sulphate solution. Calcined material is added to the acid solution until this pH value is reached. The leaching temperature lies between 50 and 70~C.
Concentrating the sludge water produces a residue of about 490 kg per ton of calcined material, i.e., 49%. The composi-tion of this residue (in relation to dry substance) is as follows:
ZnO = 14.8% CuO =0.71%
ZnS = 3.4% CdO =0.09%
ZnSO4 = 6.2% AS23 ~0.24%
ZnO.Fe2O3 = 50.6% n2 0.38%
Fe ~OH)3 = 8.7% g2 600 g/t PbSO4 = 5.7% Remainder = about 3.20%
(The "remainder" consists of SiO2, A1203, CaSO4 MgO, etc.).
This residue is dried to a fine powder. It is mixed with 30% by weiyht of finely-yround coke breeze (about 200 mesh~
and is fed uniformly, suspended in preheated air, into a fusion cyclone.
In addition to the coke breeze mentioned above, SiO2 and FeS2 are mixed with the finely-granular leaching residue as additives, for the purpose of forming slag and rock.
Upon leaving the fusion reactor, the gases and molten products are separated. The hot gases pass to an after-burning chamber where, with the addition of air, carbon monoxide and the volatile metal vapours are reoxidized, i.e., burned. These gases pass through a cooler into a filter from which the metal-oxide dusts, which are carried along and contain large amounts of the volatile components, are recovered in the form of mixed oxides.
The molten products are collected in a settling hearth where a slag/rock separation is carried out. The rock contains non-volatile metals, e.g., copper, nickel and noble metals in the enriched state, whereas the gangue, and most of the iron, pass into the slag.
rrhe mixed oxides are the products of the following chem-ical reactions:
2 _ ~ FeS + 1/2 S2 ~-C ~ 2 ~ C2 C2 + C _ _ ~ 2co ZnO + CO ~ Zn + CO2 ZnO Fe2O3 + 2CO ~ Zn + 2 FeO ~ 2 CO2 ZnSO4 + 4CO > ZnS + 4 CO2 ZnS + FeO + CO _ _ ) Zn + FeS + CO2 PbSO4 -~ 4 CO_ > PbS + 4 CO2 CdO + CO ~ Cd + C02 After-burning of the metal vapours produces:
Zn + 1/2 2 - ~ > ZnO
Cd + 1/2 2 ~ CdO
PbS ~ 3/2 2 ~ PbO + S02 PbS + 1/2 2 > PbS -~ 2 2 ) PbS04 The composition of the mixed oxide appearing in the dust collector is as follows:
ZnO = 75.3% 5dO = 0.9%
PbO = 13.3% 2 3 = 0.8%
The remainder, 9.7% consists of impurities carried 1.0 along,for example Fe203, C, SiO2 etc.
As previously described, the mixed oxides are sub-jected to leaching and precipitation, the constituents being relatively easily separated from each other.
.
Claims (8)
1. A method for recovering zinc from oxidic zinc ores or from calcined zinc-sulphide ore concentrates, or both, the ores or concentrates being leached and leaching residues thus obtain-ed being subjected to a pyrometallurgical volatilization pro-cess, comprising the steps of neutral leaching calcined zinc material with sulphuric acid solution so as to produce a leach-ing residue having a high content of undissolved zinc; drying the leaching residue by bringing the residue into contact with hot gases under reducing conditions and in a state of suspension, volatilizing said residue at a temperature of about 1450°C and precipitating the zinc and remaining volatile metal compounds in the form of mixed oxides, whereas non-volatile metals and metal compounds occur as slag or rock; and, finally, subjecting the mixed oxides to neutral leaching with a weak sulphuric acid solution.
2. A method according to claim 1, wherein neutral leach-ing of the calcined zinc material is carried out in a single stage.
3. A method according to claim 2, wherein neutralizing of the neutral leach is continued to a degree of acidity corres-ponding to a pH value of about 5, until iron which has passed into solution is precipitated, and until metal ions of other accompanying metals, which passed into solution with the floccu-lating iron compounds, are also precipitated.
4. A method according to claim 2, wherein neutralizing of the neutral leach is continued to a degree of acidity corres-ponding to a pH value of 5.5.
5. A method according to claims 1, 2 or 3, wherein pyro-metallurgical treatment of the leaching residue takes place in a fusion cyclone.
6. A method according to claim 1, wherein the mixed oxides obtained by pyrometallurgy are subjected, separately from the calcined zinc material, to neutral leaching in a weak sulphuric acid solution, a main part of the zinc oxide, contained in easily soluble form in the mixed oxides, passing into solution by preference, while the resulting leaching residue is enriched with remaining valuable metal substances.
7. A method according to claim 6, wherein with separate leaching and precipitation of the mixed oxides, this leaching and precipitation proceeds according to the leaching and pre-cipitation capabilities of the different substances of value, in such a manner that there is consecutive enrichment of those substances of value that are difficult to dissolve.
8. A method according to claims 1, 2 or 3, wherein the mixed oxides are used, during the primary leaching of the cal-cined zinc material, towards the end of the neutralizing, at a pH value of at least 2.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA283,129A CA1106620A (en) | 1977-07-20 | 1977-07-20 | Method for recovering zinc |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA283,129A CA1106620A (en) | 1977-07-20 | 1977-07-20 | Method for recovering zinc |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CA1106620A true CA1106620A (en) | 1981-08-11 |
Family
ID=4109178
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CA283,129A Expired CA1106620A (en) | 1977-07-20 | 1977-07-20 | Method for recovering zinc |
Country Status (1)
| Country | Link |
|---|---|
| CA (1) | CA1106620A (en) |
Cited By (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN102899502A (en) * | 2012-10-08 | 2013-01-30 | 来宾华锡冶炼有限公司 | Method for extracting indium-zinc and recovering tin from high tin high indium-zinc leaching residues |
-
1977
- 1977-07-20 CA CA283,129A patent/CA1106620A/en not_active Expired
Cited By (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN102899502A (en) * | 2012-10-08 | 2013-01-30 | 来宾华锡冶炼有限公司 | Method for extracting indium-zinc and recovering tin from high tin high indium-zinc leaching residues |
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| US5667553A (en) | Methods for recycling electric arc furnace dust | |
| US4619814A (en) | Process for the recovery of non-ferrous metals from sulphide ores and concentrates | |
| US6699446B2 (en) | Methods for separation of titanium from ore | |
| US4415540A (en) | Recovery of non-ferrous metals by thermal treatment of solutions containing non-ferrous and iron sulphates | |
| US4610721A (en) | Two-stage leaching process for steel plant dusts | |
| CA1224926A (en) | Method for working-up complex sulphidic ore concentrates | |
| WO1998036102A1 (en) | Refining zinc sulphide ores | |
| CN106119556A (en) | A kind of Application way of steel plant zinc smoke ash | |
| KR20060026950A (en) | Recovery of non-ferrous metals from zinc residues | |
| Shamsuddin | Metal recovery from scrap and waste | |
| CN112609084A (en) | Comprehensive treatment method for smoke dust with high zinc, lead and tin contents in fuming furnace | |
| CA1125031A (en) | Process for the roasting and chlorination of finely-divided iron ores and/or concentrates containing non-ferrous metals | |
| US3961941A (en) | Method of producing metallic lead and silver from their sulfides | |
| CA1160055A (en) | Method for the recovery of valuable metals from finely-divided pyrite | |
| US5585079A (en) | Method for leaching material containing zinc oxide and zinc silicate | |
| US4224122A (en) | Recovery of non-ferrous metals by thermal treatment of solutions containing non-ferrous and iron sulphates | |
| US4292283A (en) | Method for the recovery of zinc | |
| CA1106620A (en) | Method for recovering zinc | |
| CA1338370C (en) | Process for recovering non-ferrous metal values, especially nickel, cobalt, copper and zinc, by using melt and melt coating sulphation, from raw materials containing said metals | |
| KR101763549B1 (en) | Method and arrangement of separating arsenic from starting materials | |
| CA1177257A (en) | Method for processing sulphidic zinc ores | |
| Mirzanova et al. | Technology for processing industrial waste containing non-ferrous metals | |
| US4642133A (en) | Process for chlorinating volatilization of metals which are present in oxidic iron ores or concentrates | |
| US4317803A (en) | Recovery of non-ferrous metals by thermal treatments of solutions containing non-ferrous and iron sulphates | |
| JPH05171310A (en) | Zinc smelting method by reducing roasting |
Legal Events
| Date | Code | Title | Description |
|---|---|---|---|
| MKEX | Expiry |