CA1214045A - Process for the recovery of indium and tin - Google Patents
Process for the recovery of indium and tinInfo
- Publication number
- CA1214045A CA1214045A CA000428089A CA428089A CA1214045A CA 1214045 A CA1214045 A CA 1214045A CA 000428089 A CA000428089 A CA 000428089A CA 428089 A CA428089 A CA 428089A CA 1214045 A CA1214045 A CA 1214045A
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- Canada
- Prior art keywords
- leach
- caustic
- solution
- tin
- indium
- 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.)
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Classifications
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B25/00—Obtaining tin
- C22B25/04—Obtaining tin by wet processes
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01G—COMPOUNDS CONTAINING METALS NOT COVERED BY SUBCLASSES C01D OR C01F
- C01G19/00—Compounds of tin
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B25/00—Obtaining tin
- C22B25/06—Obtaining tin from scrap, especially tin scrap
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B58/00—Obtaining gallium or indium
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/02—Working-up flue dust
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
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- Chemical & Material Sciences (AREA)
- Engineering & Computer Science (AREA)
- Organic Chemistry (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Metallurgy (AREA)
- Manufacturing & Machinery (AREA)
- Inorganic Chemistry (AREA)
- Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Geology (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
ABSTRACT
A process for the separate recovery of indium and tin from indium-and tin-bearing materials chosen from fumes, electrolytic slimes and flue dusts which process comprises the steps of leaching said materials in oxidic or sulfatic form with a solution of caustic in a caustic leach; subjecting the leach residue to a fusion with solid caustic; leaching the fusion product with water; recovering the water leach residue as an indium concentrate; treating the water leach solution with calcium hydroxide; and recovering a concentrate containing tin as calcium hexa-hydroxo stannate. This process has the advan-tage of being simpler and yielding higher recoveries of indium and tin than most processes proposed hithero. Materials not in oxidic or sulfatic form may be converted by a fuming operation.
A process for the separate recovery of indium and tin from indium-and tin-bearing materials chosen from fumes, electrolytic slimes and flue dusts which process comprises the steps of leaching said materials in oxidic or sulfatic form with a solution of caustic in a caustic leach; subjecting the leach residue to a fusion with solid caustic; leaching the fusion product with water; recovering the water leach residue as an indium concentrate; treating the water leach solution with calcium hydroxide; and recovering a concentrate containing tin as calcium hexa-hydroxo stannate. This process has the advan-tage of being simpler and yielding higher recoveries of indium and tin than most processes proposed hithero. Materials not in oxidic or sulfatic form may be converted by a fuming operation.
Description
This invention relates to a process for the recovery of indium and tin and, more particularly, to a metallurgical process -for separately recovering indium and tin from metallurgical intermediate products.
ln the processing of complex sulfide ores and concentrates for the recovery of primary metals such as lead, zinc and copper, intermediate products are formed which may contain indium and tin, as well as many other elements such as zinc, lead, copper, iron, arsenic, antimony, cadmium, bismuth, silver and gold. Such intermediate products include flotation concentrates and tailings, slags, fumes, drosses, residues and electrolytic slimes. Where indium and tin values are sufficient to warrant recovery, a recovery process demands the separation of these metals from the other metals, either dircctly or from secondary concentrates.
The recovery of indium and tin from intermediate products obtained in the processing of complex lead-zinc concentrates has heretofore been accom-plished by any of a number of methods alone and in combination, such as chloridizing or sulfating, roasting, fuming, acid leaching, caustic fusion, precipitation as phosphate or stannate, and elec-trolysis. According to United States patent 2,052,387, which issued August 25, 1936, indium-bearing material is subjected to roasting or blast furnace treatment and the resulting indium-containing product is leached with sulfuric acidO 'I'he indium is precipitated as indium hydroxide by neutralization of the resulting solution with zinc oxide. The indium precipitate is rcdissolved and the indium is reprecipitated as indium sulfide. Impurities such as copper, arsenic, antimony and tin co-precipitate with the sulfide and can be removed from the sulfide by leaching the precipitate with sodium hydroxide.
According to United States patent 2,12~,180, which issued July 19, 1938, metallic tin is removed from lead alloys by treatment with molten caustic soda and the spent reagent containing sodium stannate is reacted with calcium - 1 - ~' s hydroxide at room temperature to form insoluble calcium stannate.
According to United States patent 2,241,438, which issued May 13, 1941, lead sulfate residue containing indium is leached with acid~ indium in the leach solution is precipitated as indium phosphate and the phosphate is converted to the hydroxide, which is roasted and reduced with hydrogen to yield indium metal.
In United States patent 2,384,610, which issued September 11, 1942, there is disclosed a method for recovering indium lrom material containing zinc, iron, aluminum and arsenic. The material is leached with dilute sulfuric acid to dissolve zinc, the leach residue is leached in strong acid, indium is precipitated, the precipitate is treated with alkali to dissolve aluminum, the residue is dissolved in sulfuric acid, indium is sponged with zinc, the sponge is dissolved and the solution is electrolyzed to recover indium.
In United States patent 2,586,649, which issued February 19, 1952, there is disclosed a methGd for the recovery of indium from an acidic solution containing indium, arsenic, cadmium, zincJ iron and lead by precipitating indium as indium arsenate, converting the arsenate to hydroxide with caustic, dissolving the hydroxide in acid and recovering indium, or igniting the hydroxide to form indium oxide. These prior art processes have the disadvan-tage of being complex and having many steps involving a number of intermediate compounds beEore the indium and/or tin are rccovered.
It has now becn ~ound that inclium and tin can be readily and sepa-rately recovered in a concentrated Eorm from metallurgical oxiclic and sulfatic intermediate products by subjecting material to treatments with sodi~ml hydroxide (caustic soda) or potassium hydroxide (caustic potash). More particularly, it has been found that by subjecting intermediate products to a leach with sodium hydroxide or potassium hydroxide solu-tion and subjecting the Leach residue to a fusion with caustic soda or caustic potash, indium and tin can be eEfectively L4~4S
separated from each other and from metals such as lead, zinc and arsenic. The process of the invention can be carried out by using either sodium hydroxide or potassium hydroxide, and with equal results. The use of sodium hydroxide is preerred for economic reasons. The two hydroxides will be re-ferred to hereinafter as caustic.
Accordingly, there is provided a process for the separate recovery of indium and tin from indium- and tin-bearing materials in oxidic or sulEatic form and including fumes and flue dusts which process comprises the steps oE
leaching said materials with a solution of caustic; subjecting th0 leach residue to a fusion with solid caustic; leaching the fusion product with water; recovering the water-leach residue as an indium concentrate; treating the water-leach solution with calcium hydroxide and recovering a concentrate containing tin as calciunt hexa-hydroxo stannate. According to a second embodi-ment of the process of the invention, there is provided a process for separate recovery of indium and tin from metallurgical intermediate products containing tin, indium, lead, ~inc, copper, iron, antimony, bismuth and arsenic which process comprises the steps of subjecting said products to a fuming operation to obtain a -fume or flue dust; leaching said fume or flue dust in a caustic leach with a solution of caustic to dissolve at least a portion of the lead, zinc and arsenic in a leach solution and to provide a leach residue containing tin and indium; subjccting said leach residue to a fusion with solid caustic to convcrt said tin to a soluble stannate; leaching the fusion product with water to substantially dissolve said solubLe stannate and any residual lead, zinc and antimony in a water lea~h solution leaving an insoluble indium containing concentrate; separating the indium-containing concentrate from said water leach solution; adding calcium hydroxide to said water leach solution to substantially precipitate said tin as calcium hexa-hydroxo stannate; recovering the precipitated tin from the residual solution as calcium hexa-hydroxo stan-nate; and returning residual solution to said caustic leach.
Preferably, the fuming operation is carried out at about lO00 to 1~00C in the presence of a reductant and an added sul:Eide; the caustic leach is carried out at a temperature in the range of from about 15C to the boiling point of the solution in a two-stage countercurrent leach with a caustic solution containing about 200 g/L caustic; the fusion is carried out at a temperature in the range ot about ~00 to 800C with an amoun-t oE solid caustic about two to three times the amount of leach residue; the leaching with water is carried out at below about 80C; and the water leach solution is treated at about 50 C with lime slurry to substantially quantitatively precipitate tin as calcium hexa-hydroxo stannate.
The invention will now be described in detail with reference to the accompanying flow sheet.
Metallurgical intermediate products that can be treated in the process of the present invention are flotation concentrates, flotation tailings, fumes, flue dusts, slags, drosses, leach residues and electrolytic slimes which are obtained in the processing of ores and concentrates for the recovery of metals such as zinc, lead and copper. Fumes, leach residues and flue dusts can usually be treated directly in a caustic leach, generally indicated at 1 while flotation concentrates and tailings, slags, drosses and electrolytic slimes must be converted into an oxidic or sulfatic Eorm whlch is suitable Eor leaching with caust:ic. Conversion may be accomplished by a pyrometalLurgical treatment such as a :Euming, oxidation, or sulEating operation~ whereby at least a portion oE the metals containcd in the Elotation concentrates and tailings, slags, drosses, and electrolytic slimes become available in a caustic leachable Eorm, such as an oxicle fume or Elue dust. For example, when drosses and slags are submitted to a fuming operation 2, shown with interrupted lines, in a ~urnace or converter, at temperatures in the range oE about 1000 to ~140~5 1400 C, major portions of the indium and the tin report in the fume emanating from the furnace or converter. Major portions of other metals such as lead, zinc, antimony, arsenic and cadmium are also volatilized and report in the fume. The fuming operation is preferably carried out at temperatures in the range of about 1000 to 1~00C and in the presence of an added recluctant such as, for example, carbon monoxide, hydrogen, coal, or coke. The addition of a sulfide, such as for example pyrrhotite, enhances the volatilization of indium and particularly that of tin. The fumed slag or dross, which may contain copper, iron, silica and lime, is removed from the furnace.
Fumes, flue dusts and leach residues are fed and subjected to a caustic leach, generally indicated at 1, with a solution of caustic. In the caustic leach 1, metals such as lead, zinc and arsenic will dissolve as metal-lates and form a leach solution, while indium and tin as well as antimony remain mainly in the leach residue. The leach is carried out at temperatures in the range of from about 15C to the boiling point of the solution at atmos-pheric pressure and preferably at a temperature of about 95C. The caustic solution which is added to the leach, and which may be a recycled solution, may contain caustic in the range of about 50 to 200 g/L. The weight ratio between the amounts of caustic and the fumes, flue dusts and residues will vary depending on the composition of the Eeed to the leach and is typically about 2:1. A retention time in the range oE about. one to Eour hours is usually sufficient Eor the completion of the leach.
The caustlc leach may be carried out in one or more stages. The leach is preferably carried out as a two-stage countercurrent leach as shown in the flowsheet. In this countercurrent leach, fume, residue or flue dust is reacted in the first stage caustic leach 3 witll caustic solution from the liquid-solids separation 6, to be described, following the second stage caustic leach 5. Additional caustic solution from liquid-solids separation 11, to be 4~
described, may also be added as indicated by the broken line. The reaction mixture from the first stage leach is subjected to a liquid-solids separation . The liquid fraction is a caustic solution which contains at leas~ a portion of the arsenic, lead, zinc and traces of tin contained in the feed to the leach. This solution is removed from the process and may be treated further for the recovery of metal values. The first leach residue from separation is subjected to the second stage caustic leach 5. Caustic in the form of recycled residual caustic solution is added. To make up the desired caustic to solids ratio in the leach, additional caustic may be added, either to the first leach 3 or the second leach 5, as desired. The reaction mixture from the second stage leach 5 is subjected to a liquid-solids separation 6. The liquid fraction from separation 6 is passed to the first stage leach 3 and the solids fraction is passed to caustic fusion 7. The first stage of the counter-current leach is preferably carried out at temperatures between about 15 C and the boiling point and the second stage at between about 80C and the boiling point of the solution at atmospheric pressure. Most preferably, both stages are conducted at a temperature of about 95C, and for a period of time of about one to two hours for each stage.
In caustic fusion 7, the solids residue from separation 6 is fused with solid caustic at a temperature in the range of about ~00 to 800C, typi-cally about 650C. The f-usion results in the formation of a soluble stannate, i.e. potassium or sodium stannate, whlch is dissoLved in a subsequcllt leach.
Prior to the Fusion, the tin (which is mainly oxicle) remains refractory and for thc most part insoluble in caustic solution. The amount of solid caustic used in the fusion is determined by the composition of the solids from sepa-ration 6, especiaLly of course the amount of tin, and is typically about two to three times the amount by weight of the solids residue. This amount of caustic represents the major input of caustic into the process. This caustic passes through the remainder of the process and is ultimately recycled to the caustic leach 1. The balance of caustic in the process is maintained by adding additional caustic directly to the caustic leach as described. The fusion requires that the material be maintained in the temperature range specified above for the period of a~out one to two hours.
The fused material Erom fusion 7 is subjected to a water leach 8 to dissolve the soluble stannate. Because the solubility of the stannate decreases with increasing caustic concentration in the leach solution and with increasing temperature, the water leach is advantageously carried out at temperatures below about 80C, pret'erably below about 50C. Any remaining lead and zinc are also dissolved in leach 8. A retention time in the range of about one to two hours is normally sufficient to complete the water leach. The leach mixture is subjected to a liquid-solids separation 9.
The solids fraction from separation 9 is a concentrate containing the indium. The indium in this concentrate represents normally at least 90%
; of the indium entering the process. The major associated impurity is antimony as the antimonate. This concentrate may be further treated by any one of a number of known methods such as leaching with acid and an oxidant followed by precipitation of indium as sulfide, or by solvent extraction of indium.
The leach solution from separation 9 is treated for the recovery of tin in tin precipitation 10. Treatment of the solution at a temperature of about 50C with calcium hydroxide, containing a slight excess, for example a 10-30% mo]ar excess, Oe calcium with respect to tin in the solution~ results in the precipitation Oe calcium hexa-hydroxo stannate. This precipitation is both selective and substantially qucultitative. 'I'he precipitate is separated from solution in liquid-solids separation 11. The solids Eraction is CaSn(OH)6, which can be readily treated Eor the recovery of tin. The solution is a caustic solution which is recycled to the caustic leach 1, or in the case oE a :~2~4(~
two-stage countercurrent leach, to the second caustic leach 5. If desired, a portion of the caustic solution recycled to leach 5 may be fed to the first caustic leach 3. The calcium stannate, which usually contains less than 1%
each of Pb, Zn, As and Sb is recovered. The overall recovery of tin is usually in the range of about 80 to 90%.
The liquid-solids separations in the process may be carried out by any of a number of known procedures, such as settling, filtrating and centri-fuging. Separation of solids from liquid is normally readily accomplished by settling and/or filtration.
The invention will now be illustrated by means of the following non-limitative examples.
Example 1 This example shows that a slag from retreatment of lead dross can be fumed to provide a material amenable to treatment for the recovery of indium and tin.
Three 50 g portions of slag were each placed in a crucible along with 5, 10 and 15 g of iron sulfide concentrate ~mostly pyrrhotite), respec-tively. Each charge was placed in a muffle furnace and a cap assembly was positioned. An inert atmosphere was maintained by admitting nitrogen gas at a rate of one L/min during the heating period. The reaction was carried out under reducing conditions at 1260C by replacing the Elow of nitrogen with carbon monoxide for 100 minutes. At the end of the period, the gas flow was turned off and; a~ter cooling, the entire crucible cmd reactlon products were ground to a fine homogeneous mixture for analyses.
The copper, silica and lime remained substantially in the residual slag, while the zlnc, arsenic and antimony, tinl and lead reported substan-tially to the fume. The indiwn divides between the slag remaining and the fume.
~L2~ S
The initial slag composition was determined. The residua]. slag and the fume were ana:Lysed for tin, indium and lead, and the distributions between the residual slag and the fume determined. The analyses are given in Table IA, and the deportments in Table IB.
TABLE IA
Slag Composition In Percentages Sn In Yb Zn Fe Cu Sb As Si.O CaO
8.7 2.4 33.8 4.3 10.5 8.4 1.6 1.6 9.4 1.1 FeS* F:inal Weight Fume Residual Slag Fume Analyses Test Slag Cor.c. Weight Loss Weight Analyses in % in %
No. in g in g :in g in % in g Sn In Pb Sn In Pb 1 50 5 33.2 39.6 21.8 5.13 2.14 15.00 12.2n 2.20 5~.70
ln the processing of complex sulfide ores and concentrates for the recovery of primary metals such as lead, zinc and copper, intermediate products are formed which may contain indium and tin, as well as many other elements such as zinc, lead, copper, iron, arsenic, antimony, cadmium, bismuth, silver and gold. Such intermediate products include flotation concentrates and tailings, slags, fumes, drosses, residues and electrolytic slimes. Where indium and tin values are sufficient to warrant recovery, a recovery process demands the separation of these metals from the other metals, either dircctly or from secondary concentrates.
The recovery of indium and tin from intermediate products obtained in the processing of complex lead-zinc concentrates has heretofore been accom-plished by any of a number of methods alone and in combination, such as chloridizing or sulfating, roasting, fuming, acid leaching, caustic fusion, precipitation as phosphate or stannate, and elec-trolysis. According to United States patent 2,052,387, which issued August 25, 1936, indium-bearing material is subjected to roasting or blast furnace treatment and the resulting indium-containing product is leached with sulfuric acidO 'I'he indium is precipitated as indium hydroxide by neutralization of the resulting solution with zinc oxide. The indium precipitate is rcdissolved and the indium is reprecipitated as indium sulfide. Impurities such as copper, arsenic, antimony and tin co-precipitate with the sulfide and can be removed from the sulfide by leaching the precipitate with sodium hydroxide.
According to United States patent 2,12~,180, which issued July 19, 1938, metallic tin is removed from lead alloys by treatment with molten caustic soda and the spent reagent containing sodium stannate is reacted with calcium - 1 - ~' s hydroxide at room temperature to form insoluble calcium stannate.
According to United States patent 2,241,438, which issued May 13, 1941, lead sulfate residue containing indium is leached with acid~ indium in the leach solution is precipitated as indium phosphate and the phosphate is converted to the hydroxide, which is roasted and reduced with hydrogen to yield indium metal.
In United States patent 2,384,610, which issued September 11, 1942, there is disclosed a method for recovering indium lrom material containing zinc, iron, aluminum and arsenic. The material is leached with dilute sulfuric acid to dissolve zinc, the leach residue is leached in strong acid, indium is precipitated, the precipitate is treated with alkali to dissolve aluminum, the residue is dissolved in sulfuric acid, indium is sponged with zinc, the sponge is dissolved and the solution is electrolyzed to recover indium.
In United States patent 2,586,649, which issued February 19, 1952, there is disclosed a methGd for the recovery of indium from an acidic solution containing indium, arsenic, cadmium, zincJ iron and lead by precipitating indium as indium arsenate, converting the arsenate to hydroxide with caustic, dissolving the hydroxide in acid and recovering indium, or igniting the hydroxide to form indium oxide. These prior art processes have the disadvan-tage of being complex and having many steps involving a number of intermediate compounds beEore the indium and/or tin are rccovered.
It has now becn ~ound that inclium and tin can be readily and sepa-rately recovered in a concentrated Eorm from metallurgical oxiclic and sulfatic intermediate products by subjecting material to treatments with sodi~ml hydroxide (caustic soda) or potassium hydroxide (caustic potash). More particularly, it has been found that by subjecting intermediate products to a leach with sodium hydroxide or potassium hydroxide solu-tion and subjecting the Leach residue to a fusion with caustic soda or caustic potash, indium and tin can be eEfectively L4~4S
separated from each other and from metals such as lead, zinc and arsenic. The process of the invention can be carried out by using either sodium hydroxide or potassium hydroxide, and with equal results. The use of sodium hydroxide is preerred for economic reasons. The two hydroxides will be re-ferred to hereinafter as caustic.
Accordingly, there is provided a process for the separate recovery of indium and tin from indium- and tin-bearing materials in oxidic or sulEatic form and including fumes and flue dusts which process comprises the steps oE
leaching said materials with a solution of caustic; subjecting th0 leach residue to a fusion with solid caustic; leaching the fusion product with water; recovering the water-leach residue as an indium concentrate; treating the water-leach solution with calcium hydroxide and recovering a concentrate containing tin as calciunt hexa-hydroxo stannate. According to a second embodi-ment of the process of the invention, there is provided a process for separate recovery of indium and tin from metallurgical intermediate products containing tin, indium, lead, ~inc, copper, iron, antimony, bismuth and arsenic which process comprises the steps of subjecting said products to a fuming operation to obtain a -fume or flue dust; leaching said fume or flue dust in a caustic leach with a solution of caustic to dissolve at least a portion of the lead, zinc and arsenic in a leach solution and to provide a leach residue containing tin and indium; subjccting said leach residue to a fusion with solid caustic to convcrt said tin to a soluble stannate; leaching the fusion product with water to substantially dissolve said solubLe stannate and any residual lead, zinc and antimony in a water lea~h solution leaving an insoluble indium containing concentrate; separating the indium-containing concentrate from said water leach solution; adding calcium hydroxide to said water leach solution to substantially precipitate said tin as calcium hexa-hydroxo stannate; recovering the precipitated tin from the residual solution as calcium hexa-hydroxo stan-nate; and returning residual solution to said caustic leach.
Preferably, the fuming operation is carried out at about lO00 to 1~00C in the presence of a reductant and an added sul:Eide; the caustic leach is carried out at a temperature in the range of from about 15C to the boiling point of the solution in a two-stage countercurrent leach with a caustic solution containing about 200 g/L caustic; the fusion is carried out at a temperature in the range ot about ~00 to 800C with an amoun-t oE solid caustic about two to three times the amount of leach residue; the leaching with water is carried out at below about 80C; and the water leach solution is treated at about 50 C with lime slurry to substantially quantitatively precipitate tin as calcium hexa-hydroxo stannate.
The invention will now be described in detail with reference to the accompanying flow sheet.
Metallurgical intermediate products that can be treated in the process of the present invention are flotation concentrates, flotation tailings, fumes, flue dusts, slags, drosses, leach residues and electrolytic slimes which are obtained in the processing of ores and concentrates for the recovery of metals such as zinc, lead and copper. Fumes, leach residues and flue dusts can usually be treated directly in a caustic leach, generally indicated at 1 while flotation concentrates and tailings, slags, drosses and electrolytic slimes must be converted into an oxidic or sulfatic Eorm whlch is suitable Eor leaching with caust:ic. Conversion may be accomplished by a pyrometalLurgical treatment such as a :Euming, oxidation, or sulEating operation~ whereby at least a portion oE the metals containcd in the Elotation concentrates and tailings, slags, drosses, and electrolytic slimes become available in a caustic leachable Eorm, such as an oxicle fume or Elue dust. For example, when drosses and slags are submitted to a fuming operation 2, shown with interrupted lines, in a ~urnace or converter, at temperatures in the range oE about 1000 to ~140~5 1400 C, major portions of the indium and the tin report in the fume emanating from the furnace or converter. Major portions of other metals such as lead, zinc, antimony, arsenic and cadmium are also volatilized and report in the fume. The fuming operation is preferably carried out at temperatures in the range of about 1000 to 1~00C and in the presence of an added recluctant such as, for example, carbon monoxide, hydrogen, coal, or coke. The addition of a sulfide, such as for example pyrrhotite, enhances the volatilization of indium and particularly that of tin. The fumed slag or dross, which may contain copper, iron, silica and lime, is removed from the furnace.
Fumes, flue dusts and leach residues are fed and subjected to a caustic leach, generally indicated at 1, with a solution of caustic. In the caustic leach 1, metals such as lead, zinc and arsenic will dissolve as metal-lates and form a leach solution, while indium and tin as well as antimony remain mainly in the leach residue. The leach is carried out at temperatures in the range of from about 15C to the boiling point of the solution at atmos-pheric pressure and preferably at a temperature of about 95C. The caustic solution which is added to the leach, and which may be a recycled solution, may contain caustic in the range of about 50 to 200 g/L. The weight ratio between the amounts of caustic and the fumes, flue dusts and residues will vary depending on the composition of the Eeed to the leach and is typically about 2:1. A retention time in the range oE about. one to Eour hours is usually sufficient Eor the completion of the leach.
The caustlc leach may be carried out in one or more stages. The leach is preferably carried out as a two-stage countercurrent leach as shown in the flowsheet. In this countercurrent leach, fume, residue or flue dust is reacted in the first stage caustic leach 3 witll caustic solution from the liquid-solids separation 6, to be described, following the second stage caustic leach 5. Additional caustic solution from liquid-solids separation 11, to be 4~
described, may also be added as indicated by the broken line. The reaction mixture from the first stage leach is subjected to a liquid-solids separation . The liquid fraction is a caustic solution which contains at leas~ a portion of the arsenic, lead, zinc and traces of tin contained in the feed to the leach. This solution is removed from the process and may be treated further for the recovery of metal values. The first leach residue from separation is subjected to the second stage caustic leach 5. Caustic in the form of recycled residual caustic solution is added. To make up the desired caustic to solids ratio in the leach, additional caustic may be added, either to the first leach 3 or the second leach 5, as desired. The reaction mixture from the second stage leach 5 is subjected to a liquid-solids separation 6. The liquid fraction from separation 6 is passed to the first stage leach 3 and the solids fraction is passed to caustic fusion 7. The first stage of the counter-current leach is preferably carried out at temperatures between about 15 C and the boiling point and the second stage at between about 80C and the boiling point of the solution at atmospheric pressure. Most preferably, both stages are conducted at a temperature of about 95C, and for a period of time of about one to two hours for each stage.
In caustic fusion 7, the solids residue from separation 6 is fused with solid caustic at a temperature in the range of about ~00 to 800C, typi-cally about 650C. The f-usion results in the formation of a soluble stannate, i.e. potassium or sodium stannate, whlch is dissoLved in a subsequcllt leach.
Prior to the Fusion, the tin (which is mainly oxicle) remains refractory and for thc most part insoluble in caustic solution. The amount of solid caustic used in the fusion is determined by the composition of the solids from sepa-ration 6, especiaLly of course the amount of tin, and is typically about two to three times the amount by weight of the solids residue. This amount of caustic represents the major input of caustic into the process. This caustic passes through the remainder of the process and is ultimately recycled to the caustic leach 1. The balance of caustic in the process is maintained by adding additional caustic directly to the caustic leach as described. The fusion requires that the material be maintained in the temperature range specified above for the period of a~out one to two hours.
The fused material Erom fusion 7 is subjected to a water leach 8 to dissolve the soluble stannate. Because the solubility of the stannate decreases with increasing caustic concentration in the leach solution and with increasing temperature, the water leach is advantageously carried out at temperatures below about 80C, pret'erably below about 50C. Any remaining lead and zinc are also dissolved in leach 8. A retention time in the range of about one to two hours is normally sufficient to complete the water leach. The leach mixture is subjected to a liquid-solids separation 9.
The solids fraction from separation 9 is a concentrate containing the indium. The indium in this concentrate represents normally at least 90%
; of the indium entering the process. The major associated impurity is antimony as the antimonate. This concentrate may be further treated by any one of a number of known methods such as leaching with acid and an oxidant followed by precipitation of indium as sulfide, or by solvent extraction of indium.
The leach solution from separation 9 is treated for the recovery of tin in tin precipitation 10. Treatment of the solution at a temperature of about 50C with calcium hydroxide, containing a slight excess, for example a 10-30% mo]ar excess, Oe calcium with respect to tin in the solution~ results in the precipitation Oe calcium hexa-hydroxo stannate. This precipitation is both selective and substantially qucultitative. 'I'he precipitate is separated from solution in liquid-solids separation 11. The solids Eraction is CaSn(OH)6, which can be readily treated Eor the recovery of tin. The solution is a caustic solution which is recycled to the caustic leach 1, or in the case oE a :~2~4(~
two-stage countercurrent leach, to the second caustic leach 5. If desired, a portion of the caustic solution recycled to leach 5 may be fed to the first caustic leach 3. The calcium stannate, which usually contains less than 1%
each of Pb, Zn, As and Sb is recovered. The overall recovery of tin is usually in the range of about 80 to 90%.
The liquid-solids separations in the process may be carried out by any of a number of known procedures, such as settling, filtrating and centri-fuging. Separation of solids from liquid is normally readily accomplished by settling and/or filtration.
The invention will now be illustrated by means of the following non-limitative examples.
Example 1 This example shows that a slag from retreatment of lead dross can be fumed to provide a material amenable to treatment for the recovery of indium and tin.
Three 50 g portions of slag were each placed in a crucible along with 5, 10 and 15 g of iron sulfide concentrate ~mostly pyrrhotite), respec-tively. Each charge was placed in a muffle furnace and a cap assembly was positioned. An inert atmosphere was maintained by admitting nitrogen gas at a rate of one L/min during the heating period. The reaction was carried out under reducing conditions at 1260C by replacing the Elow of nitrogen with carbon monoxide for 100 minutes. At the end of the period, the gas flow was turned off and; a~ter cooling, the entire crucible cmd reactlon products were ground to a fine homogeneous mixture for analyses.
The copper, silica and lime remained substantially in the residual slag, while the zlnc, arsenic and antimony, tinl and lead reported substan-tially to the fume. The indiwn divides between the slag remaining and the fume.
~L2~ S
The initial slag composition was determined. The residua]. slag and the fume were ana:Lysed for tin, indium and lead, and the distributions between the residual slag and the fume determined. The analyses are given in Table IA, and the deportments in Table IB.
TABLE IA
Slag Composition In Percentages Sn In Yb Zn Fe Cu Sb As Si.O CaO
8.7 2.4 33.8 4.3 10.5 8.4 1.6 1.6 9.4 1.1 FeS* F:inal Weight Fume Residual Slag Fume Analyses Test Slag Cor.c. Weight Loss Weight Analyses in % in %
No. in g in g :in g in % in g Sn In Pb Sn In Pb 1 50 5 33.2 39.6 21.8 5.13 2.14 15.00 12.2n 2.20 5~.70
2 50 10 37.2 38.1 22.8 2.48 1.68 15.90 15.00 2.50 48.10
3 50 15 38.2 41.2 26.8 2.30 1.72 11.90 12.90 2.00 46.20 * added as iron sulfide concentrate s TABLE IB
_ Deport nts Test Residual Slag Fume No. Sn In Pb Sn In Pb 1 Weight in g 1.70 0.71 4.97 2.65 0.49 11.90 Distribution %39.1 59.2 29.4 60.9 40.8 70.6 2 Weight in g 0.92 0.62 5.93 3.43 0.58 11.00 Distribution %21.1 51.7 35.1 78.9 48.3 65.0 3 Weight in g 0.88 0.66 4.53 3.47 0.54 12.40 Distribution %20.2 54.8 26.8 79.8 45.2 73.2 Example 2 Fumes similar in composition as the fume composition in Table IA
were leached with a 200 g/L sodium hydroxide solution at 95C for 4 hours.
The resulting residue analyzed 8.4% In, 22.4% Sn, 32.4% Zn and 5.2% Pb. The leach solution contained 1.4 g/L Sn, 19.5 g/L Zn and 17.5 g/L Pb. 25 g of the leach residue was fused with 65 g solid caustic at 800C for 30 minutes and the fusion product was subsequently leached with water. The final results showed that 80% of the tin, 60% of the zinc and 50% oE the lead were extracted from the fumes, leaving a residue analyzing 23.9% In, 10.4% Sn and 3.4% Sb.
The results Oe the leaching test show that fumes obtained from a slag are amenable to caustic leaching.
Example 3 This example illustrates the process of the invention by means of a four-pass loc]c test. All streams wcre circulated in the manner as shown on the flowsheet and all intermediate solutions were sampled. Caustlc leaches -- ],0 --for increments of 200 g of an electric furnace flue dust were carried out at 95 C with 2L of solution containing 200 g/L NaOH and for a 2 hour retention time. Total input to the test was 22.0 g In, 121.9 g Sn, 388.6 g Pb, 78.6 g Zn, 26.6 g As and 28.4 g Sb. The liquid/solids separation after the first caustic leach was made by settling, resulting in a supernatant liquor to settled solids ratio of 10:1. Settled slurry was used as the input to the second caustic leach. The liquid/solids separation after the second caustic leach was made by settling followed by filtration. The residue was not washed and was dried at 110C. The total amount of dried residue (66.7 g) was fused with caustic (first three passes 400 g NaOH each; fourth pass 200 g NaOH) in a 250 ml nickel crucible at 650C for 90 minutes. The fusion product was leached in 2L of water at 55C for 2h. The liquid/solids separation was by settling and filtration. The water leach residue (indium concentrate) was dried at 110C. The caustic fusion leach solution was treated with about the stoichio-metric quantity of Ca(OH)2 at 80C for 2h for the precipitation of calcium hexa-hydroxo stannate. The system was cooled and the slurry allowed to settle.
The settled solids filtered rapidly and were displacement washed, giving a residue of tin concentrate.
The analyses of the fourth pass of the caustic leaches, the water leach and the tin precipitation are given in Table IIA, and the metal distri-butions are given in Table IIB.
TABLE IIA
_. .
FOURTH PASS ANALYSES
-In Sn Pb Zn As S~
g/L g/L g/L g/L g/L g/L
~IRST CAUSTIC LEACH
sol'n. in 2000 mL 0.037 0.64 26.0 4.3 0.75 0.700 out 1680 mL (purge) 0.037 1.2 43.0 12.3 1.80 0.069 SECOND CAUSTIC LEACH
sol'n. in 1900 mL 0.008 0.33 3.7 2.2 0.11 0.033 out 2075 mL 0.035 0.47 21.0 3.1 0.56 1.350 WATER LEACH
leach res.
~indium concentrate) 29.0%4.0% 4.2%2.3% 0.004% 13.3%
sol'n. out 2100 mL 14 13.5 3.5 2.0 0.08 57 TIN PRECIPITATION
sol'n. in 2000 mL 0.014 13.5 3.5 2.0 0.08 0.057 out 2000 mL 0.006 0.1 3.9 2.1 0.08 0.018 residue ~tin concentrate) 0.09% 36.8%0.27% 0.33% 0.013% 0.8%
TABLE~IIB
WEIGHT
In Sn Pb Zn As Sb g g g g g g FIRST CAUSTIC LEACH
sol'n in 2000 mL 0.07~ 1.2852.00 8.60 1.50 1.~0 out 1680 mL tpurge) 0.0622.0272.20 20.66 3.02 0.12 sol'n gain ~loss)(0.012)0.7420.20 12.06 1.52 ~1.28) SECOND CAUSTIC LEACH
sol'n in 1900 mL 0.015 0.63 7.03 ~.18 0.21 0.06 out 2075 mL 0.073 0.98~3.58 6.43 1.16 2.80 sol'n gain 0.058 0.3536.55 2.25 0.95 2.7 WATER LEACH
leach res.
(indium concentrate) ~.93 0.68 0.71 0.39 _ 2.26 sol'n out 2100 mL0.03 28.357.35 ~.20 0.17 0.12 TIN PRECIPITATION
_ sol'n in 2000 mL 0.028 27.007.00 ~.O0 0.16 0.11 out 2000 mL 0.012 0.20 7.80 ~.20 0.16 0.04 residue ~tin concentrate) 0.043 25.250.13 0.16 0.01 0.38 BALANCE
dust input in 200 g 5.5030.~7 97.15 19.6~ 6.66 7.09 acco~mted output 5.03 27.8965.78 19.26 2.65 ~.14 Unaccounted loss ~gain) 8.6% 8.5% 32.3% 1.93% 60.2% ~1.6%
recovery 89.6% 82.9%
~Z~4Q4~
The results of the lock test indicate excellent recoveries of In in the final concentrate and of Sn as calcium hexa-hydroxo stannate. The indium recovery for the fourth pass was 89.6%. The indium concentrate assayed In 29%, Sn 4.0%, Pb 4.2%, Zn 2.3~, As 0.004%, and Sb 13.3%. The tin recovery for the fourth pass was 82.9%. The tin concentrate assayed Sn 36.8%, In 0.09%, Pb 0.27%, Zn 0.33%, As 0.013%, Sb 0.8%. The first caustic leach solution (purge for Pb, Zn, and As) assayed In 0.037 g/L, Sn 1.2 g/L, Pb 43 g/L, Zn 12.3 g/L, As 1.8 g/L and Sb 0.069 g/L.
200 g of input dust in the fourth pass yielded 17.0 g of indium concentrate and 68.6 g of tin concentrate. For this treatment, 400 g of NaOH
and 19 g of Ca~OH)2 were used.
The recoveries for indium, tin and lead for the four-pass lock test are given in Tables III, IV and V, respectively. From the inputs and reco-veries it can be calculated that the overall recovery of Indium was 80.6%, of tin 88.1%, and of lead 85.1%.
TABLE III
Indium Concentrate Weights Indium AnalysesIndium Recoveries in g in % in g 1st pass 13.4 30.6 4.10 2nd pass 23.4 18.4 4.31 3rd pass 22.0 20.0 4.40 4th pass 17.0 29.0 4.93 '['OTAI, 17.-/4 ~4~
TABL~ IV
Tin Concentrate Water-Leach Solution Tin-Precipitation Solution Precipitated Tin volume analysis Tin volume analysis TinRecovered in ml iJl g/L Sn in g in ml in g/L Sn in g in g -1st pass 1950 13.5 26.33 1850 1.20 2.22 24.11 2nd pass 1980 14.5 28,71 2000 0~31 0.62 28.09 3rd pass 2000 14.5 29.00 2000 0.33 0.66 28.34 4th pass 2000 13.5 27,00 2000 0.10 0.20 26.80 TOTAL: 107.34 TABLE V
Lead in First Caustic Leach Solution ~purge~
Volume Lead AnalysesLead in ml in g/L in g 1st pass 1810 51.0 92.3 2nd pass 1860 44.5 82.8 3rd pass 1675 18.4 30.8 4th pass 1680 43.0 72.2 Other 4th pass inteImediates 52.4 TOTAL 330.5 The very low In concentration in the solutions (Table II) indicate that recoveries Oe In and Sn in an industrial scale process are likely to be considerably higher.
It is 1mderstood that variations can be made in the process according to the invention without departing Erom the scope oE the invention. Eor example, a portion of the leach resiclue from the leach with caustic solution can be removed Erom the process and directly treated Eor the recovery oE
indium either by itself or together with the indium concentrate obtained after 4Q~5 the removal of tin. Such a procedure can be usef~ll if low concentrations of tin are present in the starting materials.
- 16 _
_ Deport nts Test Residual Slag Fume No. Sn In Pb Sn In Pb 1 Weight in g 1.70 0.71 4.97 2.65 0.49 11.90 Distribution %39.1 59.2 29.4 60.9 40.8 70.6 2 Weight in g 0.92 0.62 5.93 3.43 0.58 11.00 Distribution %21.1 51.7 35.1 78.9 48.3 65.0 3 Weight in g 0.88 0.66 4.53 3.47 0.54 12.40 Distribution %20.2 54.8 26.8 79.8 45.2 73.2 Example 2 Fumes similar in composition as the fume composition in Table IA
were leached with a 200 g/L sodium hydroxide solution at 95C for 4 hours.
The resulting residue analyzed 8.4% In, 22.4% Sn, 32.4% Zn and 5.2% Pb. The leach solution contained 1.4 g/L Sn, 19.5 g/L Zn and 17.5 g/L Pb. 25 g of the leach residue was fused with 65 g solid caustic at 800C for 30 minutes and the fusion product was subsequently leached with water. The final results showed that 80% of the tin, 60% of the zinc and 50% oE the lead were extracted from the fumes, leaving a residue analyzing 23.9% In, 10.4% Sn and 3.4% Sb.
The results Oe the leaching test show that fumes obtained from a slag are amenable to caustic leaching.
Example 3 This example illustrates the process of the invention by means of a four-pass loc]c test. All streams wcre circulated in the manner as shown on the flowsheet and all intermediate solutions were sampled. Caustlc leaches -- ],0 --for increments of 200 g of an electric furnace flue dust were carried out at 95 C with 2L of solution containing 200 g/L NaOH and for a 2 hour retention time. Total input to the test was 22.0 g In, 121.9 g Sn, 388.6 g Pb, 78.6 g Zn, 26.6 g As and 28.4 g Sb. The liquid/solids separation after the first caustic leach was made by settling, resulting in a supernatant liquor to settled solids ratio of 10:1. Settled slurry was used as the input to the second caustic leach. The liquid/solids separation after the second caustic leach was made by settling followed by filtration. The residue was not washed and was dried at 110C. The total amount of dried residue (66.7 g) was fused with caustic (first three passes 400 g NaOH each; fourth pass 200 g NaOH) in a 250 ml nickel crucible at 650C for 90 minutes. The fusion product was leached in 2L of water at 55C for 2h. The liquid/solids separation was by settling and filtration. The water leach residue (indium concentrate) was dried at 110C. The caustic fusion leach solution was treated with about the stoichio-metric quantity of Ca(OH)2 at 80C for 2h for the precipitation of calcium hexa-hydroxo stannate. The system was cooled and the slurry allowed to settle.
The settled solids filtered rapidly and were displacement washed, giving a residue of tin concentrate.
The analyses of the fourth pass of the caustic leaches, the water leach and the tin precipitation are given in Table IIA, and the metal distri-butions are given in Table IIB.
TABLE IIA
_. .
FOURTH PASS ANALYSES
-In Sn Pb Zn As S~
g/L g/L g/L g/L g/L g/L
~IRST CAUSTIC LEACH
sol'n. in 2000 mL 0.037 0.64 26.0 4.3 0.75 0.700 out 1680 mL (purge) 0.037 1.2 43.0 12.3 1.80 0.069 SECOND CAUSTIC LEACH
sol'n. in 1900 mL 0.008 0.33 3.7 2.2 0.11 0.033 out 2075 mL 0.035 0.47 21.0 3.1 0.56 1.350 WATER LEACH
leach res.
~indium concentrate) 29.0%4.0% 4.2%2.3% 0.004% 13.3%
sol'n. out 2100 mL 14 13.5 3.5 2.0 0.08 57 TIN PRECIPITATION
sol'n. in 2000 mL 0.014 13.5 3.5 2.0 0.08 0.057 out 2000 mL 0.006 0.1 3.9 2.1 0.08 0.018 residue ~tin concentrate) 0.09% 36.8%0.27% 0.33% 0.013% 0.8%
TABLE~IIB
WEIGHT
In Sn Pb Zn As Sb g g g g g g FIRST CAUSTIC LEACH
sol'n in 2000 mL 0.07~ 1.2852.00 8.60 1.50 1.~0 out 1680 mL tpurge) 0.0622.0272.20 20.66 3.02 0.12 sol'n gain ~loss)(0.012)0.7420.20 12.06 1.52 ~1.28) SECOND CAUSTIC LEACH
sol'n in 1900 mL 0.015 0.63 7.03 ~.18 0.21 0.06 out 2075 mL 0.073 0.98~3.58 6.43 1.16 2.80 sol'n gain 0.058 0.3536.55 2.25 0.95 2.7 WATER LEACH
leach res.
(indium concentrate) ~.93 0.68 0.71 0.39 _ 2.26 sol'n out 2100 mL0.03 28.357.35 ~.20 0.17 0.12 TIN PRECIPITATION
_ sol'n in 2000 mL 0.028 27.007.00 ~.O0 0.16 0.11 out 2000 mL 0.012 0.20 7.80 ~.20 0.16 0.04 residue ~tin concentrate) 0.043 25.250.13 0.16 0.01 0.38 BALANCE
dust input in 200 g 5.5030.~7 97.15 19.6~ 6.66 7.09 acco~mted output 5.03 27.8965.78 19.26 2.65 ~.14 Unaccounted loss ~gain) 8.6% 8.5% 32.3% 1.93% 60.2% ~1.6%
recovery 89.6% 82.9%
~Z~4Q4~
The results of the lock test indicate excellent recoveries of In in the final concentrate and of Sn as calcium hexa-hydroxo stannate. The indium recovery for the fourth pass was 89.6%. The indium concentrate assayed In 29%, Sn 4.0%, Pb 4.2%, Zn 2.3~, As 0.004%, and Sb 13.3%. The tin recovery for the fourth pass was 82.9%. The tin concentrate assayed Sn 36.8%, In 0.09%, Pb 0.27%, Zn 0.33%, As 0.013%, Sb 0.8%. The first caustic leach solution (purge for Pb, Zn, and As) assayed In 0.037 g/L, Sn 1.2 g/L, Pb 43 g/L, Zn 12.3 g/L, As 1.8 g/L and Sb 0.069 g/L.
200 g of input dust in the fourth pass yielded 17.0 g of indium concentrate and 68.6 g of tin concentrate. For this treatment, 400 g of NaOH
and 19 g of Ca~OH)2 were used.
The recoveries for indium, tin and lead for the four-pass lock test are given in Tables III, IV and V, respectively. From the inputs and reco-veries it can be calculated that the overall recovery of Indium was 80.6%, of tin 88.1%, and of lead 85.1%.
TABLE III
Indium Concentrate Weights Indium AnalysesIndium Recoveries in g in % in g 1st pass 13.4 30.6 4.10 2nd pass 23.4 18.4 4.31 3rd pass 22.0 20.0 4.40 4th pass 17.0 29.0 4.93 '['OTAI, 17.-/4 ~4~
TABL~ IV
Tin Concentrate Water-Leach Solution Tin-Precipitation Solution Precipitated Tin volume analysis Tin volume analysis TinRecovered in ml iJl g/L Sn in g in ml in g/L Sn in g in g -1st pass 1950 13.5 26.33 1850 1.20 2.22 24.11 2nd pass 1980 14.5 28,71 2000 0~31 0.62 28.09 3rd pass 2000 14.5 29.00 2000 0.33 0.66 28.34 4th pass 2000 13.5 27,00 2000 0.10 0.20 26.80 TOTAL: 107.34 TABLE V
Lead in First Caustic Leach Solution ~purge~
Volume Lead AnalysesLead in ml in g/L in g 1st pass 1810 51.0 92.3 2nd pass 1860 44.5 82.8 3rd pass 1675 18.4 30.8 4th pass 1680 43.0 72.2 Other 4th pass inteImediates 52.4 TOTAL 330.5 The very low In concentration in the solutions (Table II) indicate that recoveries Oe In and Sn in an industrial scale process are likely to be considerably higher.
It is 1mderstood that variations can be made in the process according to the invention without departing Erom the scope oE the invention. Eor example, a portion of the leach resiclue from the leach with caustic solution can be removed Erom the process and directly treated Eor the recovery oE
indium either by itself or together with the indium concentrate obtained after 4Q~5 the removal of tin. Such a procedure can be usef~ll if low concentrations of tin are present in the starting materials.
- 16 _
Claims (12)
PROPERTY OR PRIVILEGE IS CLAIMED ARE DEFINED AS FOLLOWS:
1. A process for the separate recovery of indium and tin from indium-and tin-bearing materials in oxidic or sulfatic form and including fumes and flue dusts which process comprises the steps of leaching said materials with a solution of caustic in a caustic leach subjecting the leach residue to a fusion with solid caustic resulting in the formation of soluble stannate; leaching the fusion product with water; recovering the water leach residue as an indium concentrate; treating the water leach solution with calcium hydroxide to preci-pitate tin in said leach solution, and recovering a concentrate containing tin as calcium hexa-hydroxo stannate.
2. A process for recovery of indium and tin from metallurgical interme-diate products containing tin, indium, lead, zinc, copper, iron, antimony, bismuth and arsenic which process comprises the steps of subjecting said products to a fuming operation to obtain a fume or flue dust; leaching said fume or flue dust in a caustic leach with a solution of caustic to dissolve at least a portion of the lead, zinc and arsenic in a leach solution and to pro-vide a leach residue containing tin and indium; subjecting said leach residue to a fusion with solid caustic to convert said tin to a soluble stannate;
leaching the fusion product with water to substantially dissolve said soluble stannate and any residual lead, zinc and antimony in a water leach solution leaving an insoluble indium-containing concentrate; separating the indium-con-taining concentrate from said water leach solution; adding calcium hydroxide to said water leach solution to substantially precipitate said tin as calcium hexa-hydroxo stannate; recovering the precipitated tin from the residual solu-tion as calcium hexa-hydroxo stannate; and returning residual solution to said caustic leach.
leaching the fusion product with water to substantially dissolve said soluble stannate and any residual lead, zinc and antimony in a water leach solution leaving an insoluble indium-containing concentrate; separating the indium-con-taining concentrate from said water leach solution; adding calcium hydroxide to said water leach solution to substantially precipitate said tin as calcium hexa-hydroxo stannate; recovering the precipitated tin from the residual solu-tion as calcium hexa-hydroxo stannate; and returning residual solution to said caustic leach.
3. A process as claimed in claim 2, wherein said metallurgical inter-mediate products are chosen from flotation concentrates, flotation tailings, electrolytic slimes, drosses and slags and wherein said fuming operation is carried out in a reducing atmosphere at temperatures in the range of about 1000 to 1400°C.
4. A process as claimed in claim 3, wherein said fuming operation is carried out at a temperature in the range of about 1000 to 1000°C in the presence of a reductant chosen from hydrogen, coal, coke and carbon monoxide.
5. A process as claimed in claim 2, 3 or 4, wherein said fuming operation is carried out in the presence of added sulfide.
6. A process as claimed in claim 1, or 2, or 3, wherein said caustic leach is carried out at a temperature in the range of from about 15°C to the boiling point of the solution at atmospheric pressure.
7. A process as claimed in claim 1, or 2, or 3, wherein said caustic leach is carried out at a temperature of about 95°C at atmospheric pressure.
8. A process as claimed in claim 1, or 2, or 3, wherein said caustic leach is carried out in a two-stage countercurrent leach at a temperature in the range of from about 15°C to the boiling point of the solution at atmos-pheric pressure, which countercurrent leach comprises the steps of feeding material chosen from fumes and flue dusts to a first stage caustic leach;
reacting the material with a solution of caustic from a second stage caustic leach; separating the first stage leach solution from the first stage leach residue; subjecting the first stage leach residue to a second stage caustic leach with residual caustic solution containing about 200 g/l, caustic; adding additional. caustic to the caustic leach when the ratio by weight of caustic to solids in the leach decreases below about 2:1; separating the second stage leach solution from the second stage leach residue; returning the separated second stage leach solution to said first stage leach; treating the second stage leach residue for the separate recovery of tin and indium and formation of residual caustic-containing solution; and recycling said residual caustic solution to said second stage caustic leach.
reacting the material with a solution of caustic from a second stage caustic leach; separating the first stage leach solution from the first stage leach residue; subjecting the first stage leach residue to a second stage caustic leach with residual caustic solution containing about 200 g/l, caustic; adding additional. caustic to the caustic leach when the ratio by weight of caustic to solids in the leach decreases below about 2:1; separating the second stage leach solution from the second stage leach residue; returning the separated second stage leach solution to said first stage leach; treating the second stage leach residue for the separate recovery of tin and indium and formation of residual caustic-containing solution; and recycling said residual caustic solution to said second stage caustic leach.
9. A process as claimed in claim 1, or 2, or 3, wherein the caustic fusion is carried out at a temperature in the range of about 400 to 800°C with an amount of solid caustic of about two to three times by weight of the amount of said leach residue.
10. A process as claimed in claim 1, or 2, or 3, wherein the leaching of the fusion product with water is carried out at a temperature below about 80°C.
11. A process as claimed in claim 1, or 23 or 3, wherein the leaching of the fusion product with water is carried out at a temperature below about 50 C.
12. A process as claimed in claim 1, or 2, or 3, wherein the water leach solution is treated at a temperature of about 50°C with calcium hydroxide containing a molar excess of about 10 to 30% calcium with respect to tin in said solution to substantially quantitatively precipitate tin as calcium hexa-hydroxo stannate.
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| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA000428089A CA1214045A (en) | 1983-05-13 | 1983-05-13 | Process for the recovery of indium and tin |
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| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA000428089A CA1214045A (en) | 1983-05-13 | 1983-05-13 | Process for the recovery of indium and tin |
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Cited By (9)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| WO1991001273A1 (en) * | 1989-07-21 | 1991-02-07 | Alcan International Limited | Method of making metal stannates |
| WO1991001272A1 (en) * | 1989-07-21 | 1991-02-07 | Alcan International Limited | Method of making alkali metal stannates |
| CN104032131A (en) * | 2013-06-27 | 2014-09-10 | 中石化上海工程有限公司 | Method for processing high-tin anode slurry |
| CN112280993A (en) * | 2020-10-30 | 2021-01-29 | 大冶市金欣环保科技有限公司 | Device and method for extracting tin from tin smelting alkaline residue by adopting water leaching and neutralization |
| CN114892008A (en) * | 2022-05-25 | 2022-08-12 | 中南大学 | Method for purifying lead electrolytic refining waste liquid of methanesulfonic acid system |
| CN114990337A (en) * | 2022-05-07 | 2022-09-02 | 江西铜业技术研究院有限公司 | Method for recovering tin in silver separating slag of copper anode slime by combining pyrogenic process and wet process |
| CN115092957A (en) * | 2022-05-16 | 2022-09-23 | 中南大学 | Sulfate acid tin dioxide composite material and preparation method thereof and method for cooperatively treating arsenic-alkali residue leaching residue through pyrometallurgy of antimony concentrate |
| CN116574908A (en) * | 2023-04-03 | 2023-08-11 | 西部矿业股份有限公司 | Process for jointly recycling zinc and indium by means of open-circuit impurity removal of electrolyte in zinc smelting process |
| CN117427450A (en) * | 2023-12-15 | 2024-01-23 | 云南乘风有色金属股份有限公司 | Dust metal recovery device of tin-smelting electric furnace |
-
1983
- 1983-05-13 CA CA000428089A patent/CA1214045A/en not_active Expired
Cited By (15)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| WO1991001272A1 (en) * | 1989-07-21 | 1991-02-07 | Alcan International Limited | Method of making alkali metal stannates |
| US5209911A (en) * | 1989-07-21 | 1993-05-11 | Alcan International Limited | Method of making metal stannates |
| US5342590A (en) * | 1989-07-21 | 1994-08-30 | Alcan International Limited | Method of making alkali metal stannates |
| WO1991001273A1 (en) * | 1989-07-21 | 1991-02-07 | Alcan International Limited | Method of making metal stannates |
| CN104032131A (en) * | 2013-06-27 | 2014-09-10 | 中石化上海工程有限公司 | Method for processing high-tin anode slurry |
| CN112280993A (en) * | 2020-10-30 | 2021-01-29 | 大冶市金欣环保科技有限公司 | Device and method for extracting tin from tin smelting alkaline residue by adopting water leaching and neutralization |
| CN114990337A (en) * | 2022-05-07 | 2022-09-02 | 江西铜业技术研究院有限公司 | Method for recovering tin in silver separating slag of copper anode slime by combining pyrogenic process and wet process |
| CN114990337B (en) * | 2022-05-07 | 2023-11-03 | 江西铜业技术研究院有限公司 | Method for recovering tin in silver separating slag of copper anode slime by combining pyrogenic process and wet process |
| CN115092957B (en) * | 2022-05-16 | 2023-08-18 | 中南大学 | Method for cooperatively disposing arsenic alkali slag leaching slag by adopting pyrometallurgy of antimony concentrate |
| CN115092957A (en) * | 2022-05-16 | 2022-09-23 | 中南大学 | Sulfate acid tin dioxide composite material and preparation method thereof and method for cooperatively treating arsenic-alkali residue leaching residue through pyrometallurgy of antimony concentrate |
| CN114892008A (en) * | 2022-05-25 | 2022-08-12 | 中南大学 | Method for purifying lead electrolytic refining waste liquid of methanesulfonic acid system |
| CN116574908A (en) * | 2023-04-03 | 2023-08-11 | 西部矿业股份有限公司 | Process for jointly recycling zinc and indium by means of open-circuit impurity removal of electrolyte in zinc smelting process |
| CN116574908B (en) * | 2023-04-03 | 2024-03-12 | 西部矿业股份有限公司 | Process for jointly recycling zinc and indium by means of open-circuit impurity removal of electrolyte in zinc smelting process |
| CN117427450A (en) * | 2023-12-15 | 2024-01-23 | 云南乘风有色金属股份有限公司 | Dust metal recovery device of tin-smelting electric furnace |
| CN117427450B (en) * | 2023-12-15 | 2024-04-12 | 云南乘风有色金属股份有限公司 | Dust metal recovery device of tin-smelting electric furnace |
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