CA1333199C - Process for the recovery of silver from the pb/ag cake - Google Patents
Process for the recovery of silver from the pb/ag cakeInfo
- Publication number
- CA1333199C CA1333199C CA000562414A CA562414A CA1333199C CA 1333199 C CA1333199 C CA 1333199C CA 000562414 A CA000562414 A CA 000562414A CA 562414 A CA562414 A CA 562414A CA 1333199 C CA1333199 C CA 1333199C
- Authority
- CA
- Canada
- Prior art keywords
- flotation
- silver
- sulfuric acid
- cake
- slurry
- 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 - Lifetime
Links
- 229910052709 silver Inorganic materials 0.000 title claims abstract description 52
- 239000004332 silver Substances 0.000 title claims abstract description 52
- 238000000034 method Methods 0.000 title claims abstract description 37
- 238000011084 recovery Methods 0.000 title description 16
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims abstract description 68
- 238000005188 flotation Methods 0.000 claims abstract description 41
- 239000002002 slurry Substances 0.000 claims abstract description 27
- 239000003795 chemical substances by application Substances 0.000 claims abstract description 15
- 238000009291 froth flotation Methods 0.000 claims abstract description 14
- 239000008396 flotation agent Substances 0.000 claims abstract 2
- 239000007788 liquid Substances 0.000 claims description 8
- 239000000706 filtrate Substances 0.000 claims description 5
- 229910052979 sodium sulfide Inorganic materials 0.000 claims description 4
- GRVFOGOEDUUMBP-UHFFFAOYSA-N sodium sulfide (anhydrous) Chemical compound [Na+].[Na+].[S-2] GRVFOGOEDUUMBP-UHFFFAOYSA-N 0.000 claims description 4
- 239000006228 supernatant Substances 0.000 claims description 3
- 230000002378 acidificating effect Effects 0.000 abstract description 3
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 description 49
- 229940032330 sulfuric acid Drugs 0.000 description 24
- 239000012141 concentrate Substances 0.000 description 23
- 239000011701 zinc Substances 0.000 description 19
- 229910052725 zinc Inorganic materials 0.000 description 16
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 description 13
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 13
- XEEYBQQBJWHFJM-UHFFFAOYSA-N iron Substances [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 8
- 239000000203 mixture Substances 0.000 description 7
- 239000003153 chemical reaction reagent Substances 0.000 description 6
- 238000007670 refining Methods 0.000 description 6
- 230000005587 bubbling Effects 0.000 description 4
- 229910052742 iron Inorganic materials 0.000 description 4
- 239000002994 raw material Substances 0.000 description 4
- 229910052717 sulfur Inorganic materials 0.000 description 4
- 238000005406 washing Methods 0.000 description 4
- 125000000217 alkyl group Chemical group 0.000 description 3
- ZOOODBUHSVUZEM-UHFFFAOYSA-N ethoxymethanedithioic acid Chemical compound CCOC(S)=S ZOOODBUHSVUZEM-UHFFFAOYSA-N 0.000 description 3
- 239000012467 final product Substances 0.000 description 3
- 238000002386 leaching Methods 0.000 description 3
- 229910052745 lead Inorganic materials 0.000 description 3
- 239000002516 radical scavenger Substances 0.000 description 3
- 238000004064 recycling Methods 0.000 description 3
- 150000003839 salts Chemical group 0.000 description 3
- 239000012991 xanthate Substances 0.000 description 3
- NWONKYPBYAMBJT-UHFFFAOYSA-L zinc sulfate Chemical compound [Zn+2].[O-]S([O-])(=O)=O NWONKYPBYAMBJT-UHFFFAOYSA-L 0.000 description 3
- 229960001763 zinc sulfate Drugs 0.000 description 3
- 229910000368 zinc sulfate Inorganic materials 0.000 description 3
- 229910000859 α-Fe Inorganic materials 0.000 description 3
- 239000002253 acid Substances 0.000 description 2
- 235000011149 sulphuric acid Nutrition 0.000 description 2
- 239000002699 waste material Substances 0.000 description 2
- YLZOPXRUQYQQID-UHFFFAOYSA-N 3-(2,4,6,7-tetrahydrotriazolo[4,5-c]pyridin-5-yl)-1-[4-[2-[[3-(trifluoromethoxy)phenyl]methylamino]pyrimidin-5-yl]piperazin-1-yl]propan-1-one Chemical compound N1N=NC=2CN(CCC=21)CCC(=O)N1CCN(CC1)C=1C=NC(=NC=1)NCC1=CC(=CC=C1)OC(F)(F)F YLZOPXRUQYQQID-UHFFFAOYSA-N 0.000 description 1
- VEXZGXHMUGYJMC-UHFFFAOYSA-M Chloride anion Chemical compound [Cl-] VEXZGXHMUGYJMC-UHFFFAOYSA-M 0.000 description 1
- 238000003723 Smelting Methods 0.000 description 1
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 1
- 239000012190 activator Substances 0.000 description 1
- 125000003118 aryl group Chemical group 0.000 description 1
- 239000006227 byproduct Substances 0.000 description 1
- 238000004140 cleaning Methods 0.000 description 1
- 238000011109 contamination Methods 0.000 description 1
- 238000007865 diluting Methods 0.000 description 1
- 239000012535 impurity Substances 0.000 description 1
- 229910052751 metal Inorganic materials 0.000 description 1
- 238000002156 mixing Methods 0.000 description 1
- 238000006386 neutralization reaction Methods 0.000 description 1
- KIACEOHPIRTHMI-UHFFFAOYSA-N o-propan-2-yl n-ethylcarbamothioate Chemical compound CCNC(=S)OC(C)C KIACEOHPIRTHMI-UHFFFAOYSA-N 0.000 description 1
- 239000007800 oxidant agent Substances 0.000 description 1
- 238000011112 process operation Methods 0.000 description 1
- 238000000746 purification Methods 0.000 description 1
- 238000007100 recyclization reaction Methods 0.000 description 1
- 238000000926 separation method Methods 0.000 description 1
- 238000003756 stirring Methods 0.000 description 1
- 150000004763 sulfides Chemical class 0.000 description 1
- BDHFUVZGWQCTTF-UHFFFAOYSA-M sulfonate Chemical compound [O-]S(=O)=O BDHFUVZGWQCTTF-UHFFFAOYSA-M 0.000 description 1
- 239000011593 sulfur Substances 0.000 description 1
- 239000002351 wastewater Substances 0.000 description 1
Classifications
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/02—Froth-flotation processes
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/001—Flotation agents
- B03D1/002—Inorganic compounds
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2203/00—Specified materials treated by the flotation agents; Specified applications
- B03D2203/02—Ores
- B03D2203/025—Precious metal ores
Landscapes
- Chemical & Material Sciences (AREA)
- Inorganic Chemistry (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
A froth flotation process for recovering silver from the Pb/Ag cake, comprising of the flotation which is carried out using sulfuric acid as a flotation agent, or, sulfuric acid and a sulfidizing agent, in the sulfuric acid concentration of 10-70 g/l of strong acidic condition, and in the slurry concentration of 100-300 g/l.
Description
The present invention relates to a process for the recovery of silver from the Pb/Ag cake which is generated during the hydrometallurgical zinc refining process, and particularly to a selective recovery of silver from the Pb/Ag cake by a froth-flotation process using sulfuric acid as a reagent in the flotation operation, and optionally adding a sulfidizing agent thereto, to give a strong acidic condition of 10-70 g/1 concentration of sulfuric acid.
The zinc concentrate that is traded today in the international market usually contains 48-53% Zn, 30-34% S,
The zinc concentrate that is traded today in the international market usually contains 48-53% Zn, 30-34% S,
2-12% Fe, 1-3% Pb, 50-500 g/ton Ag. Among these, silver at the present price level constitutes about 5% of the value of zinc in the zinc concentrate and therefore, silver is an important by-product and when recovered in a saleable form, it greatly economizes zinc refining due to the salvaged silver.
The hydrometallurgical refining of zinc consists of: a roasting stage of concentrated ore; a leaching stage in which roasted oxide ore is leached in sulfuric acid to obtain a solution of zinc sulfate; a purification stage removing the impurity from the solution of zinc sulfate; and an electrolytic stage in which metallic zinc is obtained from a pure solution of a zinc sulfate. Silver and lead in the zinc concentrate do not dissolve in the sulfuric acid solution during the leaching stage and remain as a residue and are called a Pb/Ag cake. The composition of the Pb/Ag cake may vary depending on the composition of zinc concentrate and the conditions of the leaching stage, but generally contain 300-1,200 g/ton of silver and 15-25% of lead.
ÇB *
Such a Pb/Ag cake is used as a raw material for a lead refining furnace or a lead and zinc simultaneously refining ISP (Imperial Smelting Process) furnace, but the lead content is low and the energy value of the cake is low, therefore its independent usage is impracticable and must be blended with others in small amounts. Due to this, its consumption has been severely limited. The method of recovering silver and lead from the unmixed Pb/Ag cake alone in the electric furnace was attempted by Mitsubishi Metal Co. of Japan, but abandoned for economic reasons.
Another approach to re-treat the Pb/Ag cake to upgrade silver and lead content utilizes the approach wherein silver and lead components in the Pb/Ag cake are dissolved in a chloride-double salt form in the presence of an oxidizing agent in the chloride solution, and then silver and lead are converted into sulfides of silver or lead, or to a metallic state which is insoluble and recovered. This process requires not only large amounts of the reagent, but also causes waste liquor (water) disposal problems in the course of the recovery of silver and lead. Therefore this process is not commercially practical.
A unique and commercially successful process today is the process for enriching the content of silver and lead by froth-flotation of Pb/Ag cake. A method of concentrating silver and lead by the froth-flotation process proposed in United States Patent No. 3,968,032 issued July 6, 1976 which discloses froth-flotation of the Pb/Ag cake which does not contain ferrite and recovery of the silver concentrate containing 14,000 g of silver per ton in the first step and , . .
B;
lead concentrate containing 55% lead in the second flotation step, respectively. In the first flotation, the pH of the flotation slurry is adjusted to 2-4.5 and a frothing collector such as xanthate, dithiophosphate, or the like is added. In the second flotation, a sulfidizing agent such as sodium sulfide and xanthate or frothing collector such as dithiophosphate is used. The example of the said invention discloses the rate of silver recovery in the first flotation to be about 60%, and that of the second flotation reached 91%.
Also, Japanese Laid Open Publication No. 52-35197 published March 17, 1977 by Vielle Montagne describes the process for treating zinc ore leach residue in which Pb/Ag cake is treated in a high-acid concentration at high temperature to eliminate ferrite completely, and then subjecting it to a first froth flotation to obtain the silver concentrate with silver content of 5,700-10,500 g/ton and then to a second flotation to obtain a lead concentrate with a 56-59% of lead, respectively. In this process too, the first flotation step pH of the flotation slurry is adjusted to 1-5, and xanthate, alkyl or aryldithiophosphate, or a mixture of alkyl and aryldithiophosphate, or isopropyl ethylthiocarbamate (addition of about 10% of sulfonate) is used as a frothing collectors, and in the second flotation, alkyl or aryl complex salt of pseudo-acid (especially inorganic salts) are used as frothing collectors. An example of the said Laid Open Publication shows the silver recovery rate of the first flotation to be 80-91%, and the lead recovery rate of 94% in the second flotation of the process.
~ B
;~`
133319~
This flotation process produces a higher grade saleable silver and lead concentrate as a final product and their recovery rate is satisfactorily high but it has the following -disadvantage; to make a Pb/Ag cake which does not contain ferrite, zinc leach residue has to be leached by strong concentrations of sulfuric acid. And then, in the flotation process, the Pb/Ag cake is diluted with a large amount of water to maintain the Pb/Ag cake in the range of pH 1-5 (or 2-4.5). Also, as the first and second flotation processes use many kinds of agents such as a frothing agent, frothing collector and activator, in addition to the higher agent cost, contamination of a large amount of water which is used to adjust the pH and slurry concentration produces a large amount of waste water.
The object of the present invention is to provide a novel flotation process for recovering silver from the Pb/Ag cake, and especially, to minimize the draw backs of the aforementioned U.S. Patent No. 3,968,032 and Japanese Laid Open Publication No. 52-35197.
The features and advantages of the present invention are attained by providing a novel flotation process for recovering silver concentrate from the Pb/Ag cake generated from the hydrometallurgical refining process by conducting Pb/Ag cake froth flotation process using sulfuric acid as a reagent having a concentration of 10-70 g/l of sulfuric acid (equivalent to pH 0.9-0.1), and by optionally adding further a small amount of sulfidizing agent thereto.
The process of the present invention is .
advantageously carried out by adding to the Pb/Ag cake, a f B i recycled liquid (filter press filtrate, centrifugal separator supernatant), water and sulfuric acid so that the flotation slurry has a slurry concentration of 100-300 g/l and an acidity of 10-70 g/l (equivalent to pH 0.9-0.1) as H2SO4 concentration. The resulting slurry is sent to the froth flotation process and silver is frothed by using air, or the resulting slurry is further treated with a sulfidizing agent such as H2S, Na2S and/or NaSH, and is then sent to the flotation process to froth the silver using air. Silver (and a small amount of lead) is collected as froth and the resulting froth is washed, filter and recovered as silver concentrate. The silver content of the silver concentrate is 8,000-12,000 grams per ton of concentrate and this figure represents an 80-90~ recovery rate.
In the present invention, even though the acidity of the Pb/Ag cake falls within a certain range, there still exists, however, variations for the individual cake.
Therefore the addition of the recycling liquor and strong sulfuric acid is adjusted so that the concentration of sulfuric acid of the flotation slurry becomes to be 10-70 g/l.
The sulfuric acid actually used as a reagent consists of the sum of the sulfuric acid self-contained in Pb/Ag cake and, the amount of sulfuric acid added to the repulper and conditioner. In the instant invention, the Pb/Ag cake contains large amounts of sulfuric acid by itself and the acidity of recycling liquor itself is also high.
Hence, the amount of sulfuric acid actually added is comparatively small.
f` B ~
The present invention provides a flotation process in which the froth flotation is conducted by adding the sulfuric acid alone as a reagent, without adding a flotation and frothing agent, or adding sulfuric acid and a sulfidizing agent to the flotation slurry.
A sulfidizing agent, such as H2S, Na2S and/or NaSH
may be added to the flotation slurry in the amount of 2.0-4.0 g/l of flotation slurry. The improvement by the invention of more than 5% of recovery rate is achieved by adding the sulfidizing agent, compared with ones not adding it.
The supply of air to froth flotation process is controlled to froth all the silver content that is recoverable and generally a supply of about 700 Nm3 per ton of Pb/Ag cake is preferable.
The characteristic features of the present invention lies in a simple and convenient process operation that consists of using only sulfuric acid, or adding some sulfidizing agent thereto, and also a unique process that facilitates the recyclization of the filtrate.
The froth flotation process of the present invention may be carried out in a number of stages, if necessary, and a washing process, a solid-liquid separation process, etc. can be installed at the site where the process is effected.
Further advantages of the present invention are attained as the operable acidic range of the froth flotation operation is widened. The acidity of actual operation is flexible and lies in a conveniently available range. The Pb/Ag cake could be added directly as a raw material for ~ B
flotation operation without a water washing or neutralization. Therefore there is no need for a neutralizer or diluting water resulting in a reduction of water consumption.
Figure 1 illustrated diagrammatically the flow of the process provided by the present invention.
Referring to the drawings in greater detail, and by reference characters thereto; as shown in Figure 1, the raw material Pb/Ag cake (1) is strongly mixed with a recycling filter process filtrate (20), centrifugal separator supernatant (19), water (3) and sulfuric acid (2) in the repulper (5) to be in a slurry state. The concentration of the slurry at the point is 300-600 g/l and concentration of -~
the sulfuric acid is 50-120 g/l. This slurry flows to conditioner (6) where water, and tailings of the cleaner cell (12) is further added to make the concentration of sulfuric -acid in the range of 10-70 g/l. In cases when sulfidizing agents are needed, they are added in the range of 0.2-4.0 g/l to the slurry, and the concentration of the slurry is finally controlled to 100-300 g/l.
The Pb/Ag cake slurry finally adjusted with sulfuric acid concentration and slurry concentration pass on to the froth flotation cell (7, 8 and 12) where flotation takes place. The flotation operation is usually carried out by bubbling the air in the range of 600-800 Nm3 per ton of Pb/Ag cake. The flotation cell is not of a special design but standard flotation equipment is acceptable. To increase the recovery rate of silver as a final product, the first flotation is carried out in the rougher cell (7) and the f. B
!' tailings of the rougher cell are frothed again at scavenger cell (8). The flotated froth from the rougher cell t7) and the scavenger cell (8) are collected at the frothed tank (11) and then pass to the cleaner cell (12) for the final cleaning flotation. The froth (17) from the cleaner cell (12) is subjected to solid-liquid separating at the settler (13) and the centrifugal separator (14) to obtain silver concentrate (23) as a final product and the tailings are returned to the repulper (5) to recover the remaining silver.
The tailings of the scavenger cell (8) are collected in the tailings tank (9), solid-liquid separated in the filter press (1) and the tailings residue (22) is discarded as a waste and the filtrate is partially recycled.
The typical grade of the Pb/Ag cake which is used as a raw material for the process of the present invention contains about 3-15% Zn, 3-13% Fe, 15-35% Pb, 12-18% S, and 200-1,500 g/ton Ag.
The silver concentrate obtained after the flotation operation contains about 6,000-12,000 grams of silver per ton of concentrate and it represents a 80-90% recovery rate.
The following examples illustrate the present invention without limiting it.
Example 1.
2 Kg (dry weight) of Pb/Ag cake having a composition of 6.2% zinc, 6% iron, 20.5% lead, 15.5% sulfur and 802 g/ton of silver was mixed with 9 R f water and 350 g of sulfuric acid for 5 minutes to obtain the slurry with the concentration of 200 g/ton and acidity of about 50 g/l by means of H2SO4.
B :
g The resulting slurry was put into the laboratory flotation cell and froth floated with air bubbling for 10 minutes and the floated froth was collected. After washing the froth with water and separating solid-liquid obtained 140 g of silver concentrate, the generated flotation tailings residue was 1,820 g.
The compositions of the silver concentrate and the flotation tailing residue were as follows:
Silver concentratefloated tailings residue (Weight ratio) (Weight ratio) Zn 6.8 4.3 Fe 2.2 5.3 Pb 17 22.1 S 55 10.3 Ag9,452 (g/t) 154(g/t) The recovery rate of silver was 82.5%.
Example 2 2 Kg of the same Pb/Ag cake used in the example 1 was mixed with 9.5 1 of water and 50 g of sulfuric acid for 3 minutes to give the slurry with about 200 g/l concentration, and acidity of about 17 g/l of H2S04. Then 200 ml of 3% NaSH
was added as a sulfuring reagent and further continued the mixing for 3 minutes. The resulting slurry was put into a laboratory flotation cell and floated by bubbling the floated froth collected. After washing the froth and separating solid-liquid obtained 148 g of silver concentrate.
also, 1859 g of the floated tailing residue was generated.
B
/
The compositions of the silver concentrate and the floated tailings residue were as follows:
Silver concentratefloated tailings residue (Weight ratio) (Weight ratio) Zn 6.9 4.7 Fe 2.0 5.2 Pb 16.0 22.3 S 56.4 10.3 Ag9,429 (g/t) 112 (g/t) The recovery rate of silver was 87%.
Example 3
The hydrometallurgical refining of zinc consists of: a roasting stage of concentrated ore; a leaching stage in which roasted oxide ore is leached in sulfuric acid to obtain a solution of zinc sulfate; a purification stage removing the impurity from the solution of zinc sulfate; and an electrolytic stage in which metallic zinc is obtained from a pure solution of a zinc sulfate. Silver and lead in the zinc concentrate do not dissolve in the sulfuric acid solution during the leaching stage and remain as a residue and are called a Pb/Ag cake. The composition of the Pb/Ag cake may vary depending on the composition of zinc concentrate and the conditions of the leaching stage, but generally contain 300-1,200 g/ton of silver and 15-25% of lead.
ÇB *
Such a Pb/Ag cake is used as a raw material for a lead refining furnace or a lead and zinc simultaneously refining ISP (Imperial Smelting Process) furnace, but the lead content is low and the energy value of the cake is low, therefore its independent usage is impracticable and must be blended with others in small amounts. Due to this, its consumption has been severely limited. The method of recovering silver and lead from the unmixed Pb/Ag cake alone in the electric furnace was attempted by Mitsubishi Metal Co. of Japan, but abandoned for economic reasons.
Another approach to re-treat the Pb/Ag cake to upgrade silver and lead content utilizes the approach wherein silver and lead components in the Pb/Ag cake are dissolved in a chloride-double salt form in the presence of an oxidizing agent in the chloride solution, and then silver and lead are converted into sulfides of silver or lead, or to a metallic state which is insoluble and recovered. This process requires not only large amounts of the reagent, but also causes waste liquor (water) disposal problems in the course of the recovery of silver and lead. Therefore this process is not commercially practical.
A unique and commercially successful process today is the process for enriching the content of silver and lead by froth-flotation of Pb/Ag cake. A method of concentrating silver and lead by the froth-flotation process proposed in United States Patent No. 3,968,032 issued July 6, 1976 which discloses froth-flotation of the Pb/Ag cake which does not contain ferrite and recovery of the silver concentrate containing 14,000 g of silver per ton in the first step and , . .
B;
lead concentrate containing 55% lead in the second flotation step, respectively. In the first flotation, the pH of the flotation slurry is adjusted to 2-4.5 and a frothing collector such as xanthate, dithiophosphate, or the like is added. In the second flotation, a sulfidizing agent such as sodium sulfide and xanthate or frothing collector such as dithiophosphate is used. The example of the said invention discloses the rate of silver recovery in the first flotation to be about 60%, and that of the second flotation reached 91%.
Also, Japanese Laid Open Publication No. 52-35197 published March 17, 1977 by Vielle Montagne describes the process for treating zinc ore leach residue in which Pb/Ag cake is treated in a high-acid concentration at high temperature to eliminate ferrite completely, and then subjecting it to a first froth flotation to obtain the silver concentrate with silver content of 5,700-10,500 g/ton and then to a second flotation to obtain a lead concentrate with a 56-59% of lead, respectively. In this process too, the first flotation step pH of the flotation slurry is adjusted to 1-5, and xanthate, alkyl or aryldithiophosphate, or a mixture of alkyl and aryldithiophosphate, or isopropyl ethylthiocarbamate (addition of about 10% of sulfonate) is used as a frothing collectors, and in the second flotation, alkyl or aryl complex salt of pseudo-acid (especially inorganic salts) are used as frothing collectors. An example of the said Laid Open Publication shows the silver recovery rate of the first flotation to be 80-91%, and the lead recovery rate of 94% in the second flotation of the process.
~ B
;~`
133319~
This flotation process produces a higher grade saleable silver and lead concentrate as a final product and their recovery rate is satisfactorily high but it has the following -disadvantage; to make a Pb/Ag cake which does not contain ferrite, zinc leach residue has to be leached by strong concentrations of sulfuric acid. And then, in the flotation process, the Pb/Ag cake is diluted with a large amount of water to maintain the Pb/Ag cake in the range of pH 1-5 (or 2-4.5). Also, as the first and second flotation processes use many kinds of agents such as a frothing agent, frothing collector and activator, in addition to the higher agent cost, contamination of a large amount of water which is used to adjust the pH and slurry concentration produces a large amount of waste water.
The object of the present invention is to provide a novel flotation process for recovering silver from the Pb/Ag cake, and especially, to minimize the draw backs of the aforementioned U.S. Patent No. 3,968,032 and Japanese Laid Open Publication No. 52-35197.
The features and advantages of the present invention are attained by providing a novel flotation process for recovering silver concentrate from the Pb/Ag cake generated from the hydrometallurgical refining process by conducting Pb/Ag cake froth flotation process using sulfuric acid as a reagent having a concentration of 10-70 g/l of sulfuric acid (equivalent to pH 0.9-0.1), and by optionally adding further a small amount of sulfidizing agent thereto.
The process of the present invention is .
advantageously carried out by adding to the Pb/Ag cake, a f B i recycled liquid (filter press filtrate, centrifugal separator supernatant), water and sulfuric acid so that the flotation slurry has a slurry concentration of 100-300 g/l and an acidity of 10-70 g/l (equivalent to pH 0.9-0.1) as H2SO4 concentration. The resulting slurry is sent to the froth flotation process and silver is frothed by using air, or the resulting slurry is further treated with a sulfidizing agent such as H2S, Na2S and/or NaSH, and is then sent to the flotation process to froth the silver using air. Silver (and a small amount of lead) is collected as froth and the resulting froth is washed, filter and recovered as silver concentrate. The silver content of the silver concentrate is 8,000-12,000 grams per ton of concentrate and this figure represents an 80-90~ recovery rate.
In the present invention, even though the acidity of the Pb/Ag cake falls within a certain range, there still exists, however, variations for the individual cake.
Therefore the addition of the recycling liquor and strong sulfuric acid is adjusted so that the concentration of sulfuric acid of the flotation slurry becomes to be 10-70 g/l.
The sulfuric acid actually used as a reagent consists of the sum of the sulfuric acid self-contained in Pb/Ag cake and, the amount of sulfuric acid added to the repulper and conditioner. In the instant invention, the Pb/Ag cake contains large amounts of sulfuric acid by itself and the acidity of recycling liquor itself is also high.
Hence, the amount of sulfuric acid actually added is comparatively small.
f` B ~
The present invention provides a flotation process in which the froth flotation is conducted by adding the sulfuric acid alone as a reagent, without adding a flotation and frothing agent, or adding sulfuric acid and a sulfidizing agent to the flotation slurry.
A sulfidizing agent, such as H2S, Na2S and/or NaSH
may be added to the flotation slurry in the amount of 2.0-4.0 g/l of flotation slurry. The improvement by the invention of more than 5% of recovery rate is achieved by adding the sulfidizing agent, compared with ones not adding it.
The supply of air to froth flotation process is controlled to froth all the silver content that is recoverable and generally a supply of about 700 Nm3 per ton of Pb/Ag cake is preferable.
The characteristic features of the present invention lies in a simple and convenient process operation that consists of using only sulfuric acid, or adding some sulfidizing agent thereto, and also a unique process that facilitates the recyclization of the filtrate.
The froth flotation process of the present invention may be carried out in a number of stages, if necessary, and a washing process, a solid-liquid separation process, etc. can be installed at the site where the process is effected.
Further advantages of the present invention are attained as the operable acidic range of the froth flotation operation is widened. The acidity of actual operation is flexible and lies in a conveniently available range. The Pb/Ag cake could be added directly as a raw material for ~ B
flotation operation without a water washing or neutralization. Therefore there is no need for a neutralizer or diluting water resulting in a reduction of water consumption.
Figure 1 illustrated diagrammatically the flow of the process provided by the present invention.
Referring to the drawings in greater detail, and by reference characters thereto; as shown in Figure 1, the raw material Pb/Ag cake (1) is strongly mixed with a recycling filter process filtrate (20), centrifugal separator supernatant (19), water (3) and sulfuric acid (2) in the repulper (5) to be in a slurry state. The concentration of the slurry at the point is 300-600 g/l and concentration of -~
the sulfuric acid is 50-120 g/l. This slurry flows to conditioner (6) where water, and tailings of the cleaner cell (12) is further added to make the concentration of sulfuric -acid in the range of 10-70 g/l. In cases when sulfidizing agents are needed, they are added in the range of 0.2-4.0 g/l to the slurry, and the concentration of the slurry is finally controlled to 100-300 g/l.
The Pb/Ag cake slurry finally adjusted with sulfuric acid concentration and slurry concentration pass on to the froth flotation cell (7, 8 and 12) where flotation takes place. The flotation operation is usually carried out by bubbling the air in the range of 600-800 Nm3 per ton of Pb/Ag cake. The flotation cell is not of a special design but standard flotation equipment is acceptable. To increase the recovery rate of silver as a final product, the first flotation is carried out in the rougher cell (7) and the f. B
!' tailings of the rougher cell are frothed again at scavenger cell (8). The flotated froth from the rougher cell t7) and the scavenger cell (8) are collected at the frothed tank (11) and then pass to the cleaner cell (12) for the final cleaning flotation. The froth (17) from the cleaner cell (12) is subjected to solid-liquid separating at the settler (13) and the centrifugal separator (14) to obtain silver concentrate (23) as a final product and the tailings are returned to the repulper (5) to recover the remaining silver.
The tailings of the scavenger cell (8) are collected in the tailings tank (9), solid-liquid separated in the filter press (1) and the tailings residue (22) is discarded as a waste and the filtrate is partially recycled.
The typical grade of the Pb/Ag cake which is used as a raw material for the process of the present invention contains about 3-15% Zn, 3-13% Fe, 15-35% Pb, 12-18% S, and 200-1,500 g/ton Ag.
The silver concentrate obtained after the flotation operation contains about 6,000-12,000 grams of silver per ton of concentrate and it represents a 80-90% recovery rate.
The following examples illustrate the present invention without limiting it.
Example 1.
2 Kg (dry weight) of Pb/Ag cake having a composition of 6.2% zinc, 6% iron, 20.5% lead, 15.5% sulfur and 802 g/ton of silver was mixed with 9 R f water and 350 g of sulfuric acid for 5 minutes to obtain the slurry with the concentration of 200 g/ton and acidity of about 50 g/l by means of H2SO4.
B :
g The resulting slurry was put into the laboratory flotation cell and froth floated with air bubbling for 10 minutes and the floated froth was collected. After washing the froth with water and separating solid-liquid obtained 140 g of silver concentrate, the generated flotation tailings residue was 1,820 g.
The compositions of the silver concentrate and the flotation tailing residue were as follows:
Silver concentratefloated tailings residue (Weight ratio) (Weight ratio) Zn 6.8 4.3 Fe 2.2 5.3 Pb 17 22.1 S 55 10.3 Ag9,452 (g/t) 154(g/t) The recovery rate of silver was 82.5%.
Example 2 2 Kg of the same Pb/Ag cake used in the example 1 was mixed with 9.5 1 of water and 50 g of sulfuric acid for 3 minutes to give the slurry with about 200 g/l concentration, and acidity of about 17 g/l of H2S04. Then 200 ml of 3% NaSH
was added as a sulfuring reagent and further continued the mixing for 3 minutes. The resulting slurry was put into a laboratory flotation cell and floated by bubbling the floated froth collected. After washing the froth and separating solid-liquid obtained 148 g of silver concentrate.
also, 1859 g of the floated tailing residue was generated.
B
/
The compositions of the silver concentrate and the floated tailings residue were as follows:
Silver concentratefloated tailings residue (Weight ratio) (Weight ratio) Zn 6.9 4.7 Fe 2.0 5.2 Pb 16.0 22.3 S 56.4 10.3 Ag9,429 (g/t) 112 (g/t) The recovery rate of silver was 87%.
Example 3
3 Kg of Pb/Ag cake (dry weight) having 9.5% Zn, 8.8% Fe, 19.5% Pb, 14.2% S and 704 g/ton of Ag was mixed with 20 1 of water and 250 g of H2S04 for 5 minutes with stirring to make a slurry with a concentration of 150 g/l and acidity of about 30 g/l. The resulting slurry was put into the laboratory cell and floated by bubbling for 10 minutes. The floated froth was collected and washed and solid-liquid separated to obtain 195 g of silver concentrate. Also, 2,765 g of the floated tailing residue was produced. The compositions of the silver concentrate and the floated tailings residue were as follows:
Silver concentrate floated tailing residue (Weight ratio) (Weight ratio) Zn 10.2 Fe 6.5 8.0 Pb 15.5 19.8 S 56.9 13.9 Ag8,803 (g/t) 143 (g/t) The recovery rate of silver was 81.3%
~ B
Silver concentrate floated tailing residue (Weight ratio) (Weight ratio) Zn 10.2 Fe 6.5 8.0 Pb 15.5 19.8 S 56.9 13.9 Ag8,803 (g/t) 143 (g/t) The recovery rate of silver was 81.3%
~ B
Claims (3)
1. In a froth flotation process for recovering silver from a Pb/Ag cake, the improvement wherein the flotation is carried out in a flotation slurry having a concentration of 100-300 g/l using sulfuric acid as a flotation agent, the sulfuric acid being present in a concentration of 10-70 g/l.
2. A process according to claim 1, wherein at least one of H2S, Na2S and NaSH is added to the slurry as a sulfidizing agent in the range of 0.2-4.0 g/l.
3. A process according to claim 1, wherein the total sulfuric acid concentration is comprised of the sum of sulfuric acid originally contained in the Pb/Ag cake, sulfuric acid added to a repulper, and sulfuric acid contained in recycled liquid which is the filter press filtrate and centrifugal separator supernatant.
Priority Applications (3)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA000562414A CA1333199C (en) | 1988-03-24 | 1988-03-24 | Process for the recovery of silver from the pb/ag cake |
| AU13589/88A AU593728B2 (en) | 1988-03-24 | 1988-03-24 | A process for the recovery of silver from the Pb/Ag cake |
| JP63072171A JPH0768589B2 (en) | 1988-03-24 | 1988-03-28 | How to recover silver from lead / silver cake |
Applications Claiming Priority (3)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA000562414A CA1333199C (en) | 1988-03-24 | 1988-03-24 | Process for the recovery of silver from the pb/ag cake |
| AU13589/88A AU593728B2 (en) | 1988-03-24 | 1988-03-24 | A process for the recovery of silver from the Pb/Ag cake |
| JP63072171A JPH0768589B2 (en) | 1988-03-24 | 1988-03-28 | How to recover silver from lead / silver cake |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CA1333199C true CA1333199C (en) | 1994-11-22 |
Family
ID=27152044
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CA000562414A Expired - Lifetime CA1333199C (en) | 1988-03-24 | 1988-03-24 | Process for the recovery of silver from the pb/ag cake |
Country Status (3)
| Country | Link |
|---|---|
| JP (1) | JPH0768589B2 (en) |
| AU (1) | AU593728B2 (en) |
| CA (1) | CA1333199C (en) |
Cited By (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN109844146A (en) * | 2016-09-14 | 2019-06-04 | 奥图泰(芬兰)公司 | Method for recycling noble metal |
Families Citing this family (4)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| FI122099B (en) * | 2010-04-30 | 2011-08-31 | Outotec Oyj | A method for recovering precious metals |
| CN102371212B (en) * | 2011-10-19 | 2013-12-18 | 昆明理工大学 | Technology of enhanced-dispersion partial selective and bulk flotation of lead and zinc sulfide ores under low and high alkalinity |
| CN103286008B (en) * | 2013-05-13 | 2015-11-18 | 湖南水口山有色金属集团有限公司 | Sulfuric acid is adopted to adjust the method for silver raising recovery rate in agent raising flotation of silver concentrate |
| CN103909018A (en) * | 2014-04-18 | 2014-07-09 | 武汉工程大学 | Collophanite reverse flotation depressing agent and application thereof |
Family Cites Families (3)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| BE829988A (en) * | 1975-06-06 | 1975-10-01 | PROCESS FOR TREATMENT OF ZINC ORE LEACHING RESIDUES | |
| FI65805C (en) * | 1980-09-30 | 1984-07-10 | Outokumpu Oy | FOERFARANDE FOER AOTERVINNING AV BLY SILVER OCH GULD UR JAERNHALTIGT AVFALL FRAON EN ELEKTROLYTISK ZINKPROCESS |
| JPS5845339A (en) * | 1981-09-14 | 1983-03-16 | Toho Aen Kk | Treatment of zinc leached slag and secondarily leached residues from said slag |
-
1988
- 1988-03-24 CA CA000562414A patent/CA1333199C/en not_active Expired - Lifetime
- 1988-03-24 AU AU13589/88A patent/AU593728B2/en not_active Expired
- 1988-03-28 JP JP63072171A patent/JPH0768589B2/en not_active Expired - Lifetime
Cited By (3)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN109844146A (en) * | 2016-09-14 | 2019-06-04 | 奥图泰(芬兰)公司 | Method for recycling noble metal |
| US20190211421A1 (en) * | 2016-09-14 | 2019-07-11 | Outotec (Finland) Oy | Method for recovering precious metal |
| US10590510B2 (en) * | 2016-09-14 | 2020-03-17 | Outotec (Finland) Oy | Method for recovering precious metal |
Also Published As
| Publication number | Publication date |
|---|---|
| JPH02133531A (en) | 1990-05-22 |
| AU593728B2 (en) | 1989-09-28 |
| AU1358988A (en) | 1989-09-28 |
| JPH0768589B2 (en) | 1995-07-26 |
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