CN101451198A - Method for recovering zinc and lead from waste electrolysis anode sludge - Google Patents
Method for recovering zinc and lead from waste electrolysis anode sludge Download PDFInfo
- Publication number
- CN101451198A CN101451198A CNA2007101685612A CN200710168561A CN101451198A CN 101451198 A CN101451198 A CN 101451198A CN A2007101685612 A CNA2007101685612 A CN A2007101685612A CN 200710168561 A CN200710168561 A CN 200710168561A CN 101451198 A CN101451198 A CN 101451198A
- Authority
- CN
- China
- Prior art keywords
- lead
- acid
- zinc
- sodium hydroxide
- filtrate
- 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.)
- Granted
Links
Images
Classifications
-
- 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
Landscapes
- Manufacture And Refinement Of Metals (AREA)
- Electrolytic Production Of Metals (AREA)
Abstract
The invention discloses a method for recovering zinc and lead from waste electrolytic anode slime. The method comprises: grinding the waste electrolytic anode slime first, adding the waste electrolytic anode slime into a mixed acid solution, stirring the mixture for 0.2 to 4 hours at a temperature of between 20 and 60 DEG C, filtering the mixture, and taking filter residue for standby; and adding filtrate into an ammonia water-ammonia carbonate solution, stirring and filtering the mixture, adding sodium sulfide solid into filtrate, obtaining lead sulfide black deposit, uniformly mixing the lead sulfide deposit and solid sodium hydroxide, performing heating reaction for 40 to 50 minutes, and performing temperature reduction, waterlogging and filtration to obtain metallic lead. The method has the following advantages: firstly, the method adopts wet-method technology and has low energy consumption, low cost and good labor conditions; secondly, the method adopts a method for performing oxidation reduction and precipitation in a solution to remove impurities, and avoids the complex purifying process; and thirdly, the method simultaneously recovers the zinc and the lead in the waste slime and has high recovery rate and high metal grade.
Description
(1) technical field: the present invention relates to the recovery method of non-ferrous metal, specifically is the method that reclaims zinc lead from waste electrolysis anode sludge.
The ㈡ background technology: over nearly 20 years, because expanding economy, the non-ferrous metal consumption rises rapidly, has produced a large amount of callable non-ferrous metal waste materials, and the production of secondary nonferrous metal has simultaneously also obtained development faster.The production of secondary nonferrous metal comes into one's own day by day with its low investment, less energy-consumption, low cost, the low pollution and the better simply characteristics of technology.
According to incompletely statistics, the leaching utilising zinc containing waste residue that hydrometallurgy produces adds that the utilising zinc containing waste residue mud accumulative total that does not obtain in the industrial production utilizing reaches 1,000 ten thousand tons, and the annual industrial residue mud that produces 320,000 tons approximately is as electroplating cadmia mud, copper-lead cadmia mud etc.If these copperized lead cadmia mud are not recycled, both can produce and pollute environment, can cause a large amount of wastings of resources again.Therefore, reclaim zinc lead from copperized lead cadmia mud, preparation zinc leadwork makes solid waste resource recovery, reduces environmental pollution, seems particularly urgent.
Utilising zinc containing waste residue is because the source is different, and its composition is all different with character, and its treatment process is a lot, mainly is divided into two kinds of pyrogenic process and wet methods.Pyrometallurgy of zinc is owing to energy consumption height, pollution are big, and labor condition is poor, metal recovery rate is low, metal content is low is eliminated gradually.General zinc hydrometallurgy is made up of operations such as concentrate roasting, flue gas acid preparing, leaching, clean liquid, electrodeposition, foundings.The major advantage of zinc hydrometallurgy is that environmental health, labor condition are good, and metal recovery rate height, metal content height are easy to large-scale production.
Chinese patent CN1243167A proposed a kind of from utilising zinc containing waste residue the extracting zinc method, this method comprises preparation, sintering, distillation, the deslagging and reinforced of zinc raw material ball, trembling in the jar that sintering and distillation are evenly disposed in the sintering matrass carried out synchronously, vertically fill in the batch can and contain the zinc raw material ball, with fire coal the sintering matrass is heated to 1300 ℃ gradually, is vertically carrying out redox reaction, the vaporization of solid zinc, zinc gas liquefaction, sintering and preheating successively in the batch can.This method can become waste material the zinc that has industrial value to utilize, but has many shortcomings: (1) this method still adopts traditional thermal process, and the Heating temperature height reaches 1300 ℃, and energy consumption is big, and production cost is too high; (2) reaction unit is complicated, is unfavorable for operation; (3) metal species of this method recovery is single, has only metallic zinc.
The technological process of production of secondary lead mainly is that liquate slightly refines-electrorefining-basic refining and traditional reverberatory smelting and blast furnace smelting, also has lead ore to smelt useless assorted plumbous secondary lead flow process, the lead 2-base alloy waste material production secondary lead alloy flow process etc. of producing of collocation processing in addition.But shortcomings such as above most technology exists step complexity, production cost height, metal recovery rate is low, metal content is low.
The ㈢ summary of the invention: purpose of the present invention just provides a kind of method that reclaims zinc lead from waste electrolysis anode sludge.Present method is raw material with the waste electrolysis anode sludge, adopts leaching-electrodeposition-high temperature alkali pretreatment to reclaim metallic zinc and lead, and technical process is short, and production cost is low, is suitable for the production of secondary nonferrous metal especially.
Method of the present invention comprises the steps:
A. the leaching of waste electrolysis anode sludge
Waste electrolysis anode sludge is ground to the 150-300 order, joins in the mixed acid solution of the doubly useless mud weight of 3-10, the volumetric molar concentration of nitration mixture is the 1-2 mol, adds useless mud weight 0.01-1 oxygenant doubly again, stirs 0.2-4 hour down at 20-60 ℃, filters, and filter residue is standby; Filtrate is adjusted to pH8.0 with sodium hydroxide solution, the ammoniacal liquor-carbonic acid ammonia solution of weight such as adding and nitration mixture, and the weight ratio of ammoniacal liquor and carbonic acid ammonia is 1:1, the concentration of ammoniacal liquor-carbonic acid ammonia solution is 20%, stirs, and filters, and gets solution of zinc sulfate;
B. electrodeposition
Above-mentioned solution of zinc sulfate is concentrated to 1 mol, after carry out electrodeposition, promptly get metallic zinc;
C. high temperature alkaline purification system lead
In above-mentioned filter residue, add 30% sodium hydroxide solution, the weight ratio of filter residue and sodium hydroxide solution is 0.1-0.5, is stirred to the filtrate clarification, filters, in filtrate, add the sodium sulphite solid again, the weight ratio of sodium sulphite and filter residue is 1:1-2, gets the lead sulfide black precipitate, and lead sulfide precipitation and solid sodium hydroxide are pressed 2: 3 weight ratio mixing, placed 300-400 ℃ of reacting by heating 40-50 minutes, reduce to room temperature, water logging is filtered and is promptly got metallic lead.
Described nitration mixture is the mixture of any two kinds of acid in sulfuric acid, nitric acid, phosphoric acid, Phenylsulfonic acid, hydrochloric acid, the tosic acid, and wherein the weight ratio of two kinds of acid is 1:1.
Described oxygenant is one or both in potassium permanganate, the vitriol oil, perchloric acid, chlorine, industrial bleaching powder, hydrogen peroxide, bromine water, the persulphate, and wherein the weight ratio of two kinds of oxygenants is 1:1.
The useless mud of anode that described waste electrolysis anode sludge obtains for the electrolytic zinc concentrate.
The present invention compared with prior art has following advantage:
(1) adopt wet processing, energy consumption is low, cost is low, environmental health, labor condition are good;
(2) redox and sedimentary method are removed one step of impurity in the employing solution, have avoided complicated purification purification process, have reduced production cost;
(3) zinc and lead in the useless mud are reclaimed simultaneously, and the rate of recovery can reach more than 80%, zinc and plumbous of high grade, the grade 83% of zinc, plumbous grade 76% has improved economic benefit.
(4) description of drawings:
Fig. 1 is a process flow diagram of the present invention.
The ㈤ embodiment:
Embodiment 1:
Get the useless mud of 10 grams and be milled down to 150 orders, join in the sulfuric acid and hydrochloric acid mixed solution of 30 grams, 1 mol concentration, wherein the proportioning of sulfuric acid and hydrochloric acid is 1:1, add 0.1 gram industrial bleaching powder again, stirred 0.2 hour, filter at 20-60 ℃, collect filtrate, filter residue is standby.Filtrate is adjusted to pH8.0 with sodium hydroxide solution, add 20% the ammoniacal liquor-carbonic acid ammonia solution of 30 grams again by the preparation of equivalent ammoniacal liquor and carbonic acid ammonia, in 60 ℃ of stirrings till not producing precipitation, filter, collect filtrate and it is concentrated to 1 mol, electrodeposition promptly obtains 0.34 gram zinc, and the rate of recovery is 83%;
Adding 30% sodium hydroxide solution in above-mentioned filter residue, the weight ratio of filter residue and sodium hydroxide solution is 0.1-0.5, stirs, and filters, and adds 20 gram sodium sulphite solids again in filtrate, filters to such an extent that black lead sulfide precipitates.Lead sulfide precipitation and solid sodium hydroxide is even with the mixed of 2:3, put into crucible, place resistance furnace to heat then, under 300-400 ℃ temperature, reacted 40-50 minutes, reduce to room temperature, add 1 premium on currency and soaked 20-30 minutes, filter, it is plumbous promptly to get 0.68 gram, and the rate of recovery is 81%.
Embodiment 2
Get the useless mud of 10 grams and be milled down to 200 orders, join in the nitric acid and hydrochloric acid mixed solution of 100 grams, 1.5 mol concentration, wherein the proportioning of nitric acid and hydrochloric acid is 1:1, add 0.5 gram potassium permanganate again, stirred 2 hours, filter at 20-60 ℃, collect filtrate, filter residue is standby.Filtrate is adjusted to pH8.0 with sodium hydroxide solution, add the 20% ammoniacal liquor-carbonic acid ammonia solution of 100 grams by equivalent ammoniacal liquor and the preparation of carbonic acid ammonia, in 60 ℃ of stirrings till not producing precipitation, filter, collect filtrate and it is concentrated to 1 mol, electrodeposition promptly obtains 0.5 gram zinc, and the rate of recovery is 85%;
Adding 30% sodium hydroxide solution in above-mentioned filter residue, the weight ratio of filter residue and sodium hydroxide is 0.1-0.5, stirs, and filters, and adds 20 gram sodium sulphite solids again in filtrate, filters to such an extent that black lead sulfide precipitates.Lead sulfide precipitation and solid sodium hydroxide is even with the mixed of 2:3, put into crucible, place resistance furnace to heat then, under 300-400 ℃ temperature, reacted 40-50 minutes, reduce to room temperature, add 1 premium on currency and soaked 20-30 minutes, filter, it is plumbous promptly to get 0.74 gram, and the rate of recovery is 88%.
Embodiment 3
Get the useless mud of 10 grams and be milled down to 180 orders, join in the sulfuric acid and phosphoric acid mixing solutions of 50 grams, 2 mol concentration, wherein the proportioning of sulfuric acid and phosphoric acid is 1:1, adds 1 gram perchloric acid again, stirs 3 hours at 20-60 ℃, filters, and collects filtrate, and filter residue is standby.Filtrate is adjusted to pH8.0 with sodium hydroxide solution, add the 20% ammoniacal liquor-carbonic acid ammonia solution of 50 grams by equivalent ammoniacal liquor and the preparation of carbonic acid ammonia, in 60 ℃ of stirrings till not producing precipitation, filter, collect filtrate and it is concentrated to 1 mol, electrodeposition promptly obtains 0.377 gram zinc, and the rate of recovery is 92%;
Adding 30% sodium hydroxide solution in above-mentioned filter residue, the weight ratio of filter residue and sodium hydroxide is 0.1-0.5, stirs, and filters, and adds 20 gram sodium sulphite solids again in filtrate, filters to such an extent that black lead sulfide precipitates.Lead sulfide precipitation and solid sodium hydroxide is even with the mixed of 2:3, put into crucible, place resistance furnace to heat then, under 300-400 ℃ temperature, reacted 40-50 minutes, reduce to room temperature, add 1 premium on currency and soaked 20-30 minutes, filter, it is plumbous promptly to get 0.739 gram, and the rate of recovery is 85%.
Embodiment 4
Get the useless mud of 10 grams and be milled down to 300 orders, join in the tosic acid and vitriolic mixing solutions of 80 grams, 1 mol concentration, wherein the proportioning of sulfuric acid and toluenesulphonic acids is 1:1, add 5 gram chlorine again, stirred 3 hours, filter at 20-60 ℃, collect filtrate, filter residue is standby.Filtrate is adjusted to pH8.0 with sodium hydroxide solution, add the 20% ammoniacal liquor-carbonic acid ammonia solution of 80 grams by equivalent ammoniacal liquor and the preparation of carbonic acid ammonia, in 60 ℃ of stirrings till not producing precipitation, filter, collect filtrate and it is concentrated to 1 mol, electrodeposition promptly obtains 0.39 gram zinc, and the rate of recovery is 95%;
Adding 30% sodium hydroxide solution in above-mentioned filter residue, the weight ratio of filter residue and sodium hydroxide is 0.1-0.5, stirs, and filters, and adds 20 gram sodium sulphite solids again in filtrate, filters to such an extent that black lead sulfide precipitates.Lead sulfide precipitation and solid sodium hydroxide is even with the mixed of 2:3, put into crucible, place resistance furnace to heat then, under 300-400 ℃ temperature, reacted 40-50 minutes, reduce to room temperature, add 1 premium on currency and soaked 20-30 minutes, filter, it is plumbous promptly to get 0.748 gram, and the rate of recovery is 89%.
Embodiment 5
Get the useless mud of 10 grams and be milled down to 200 orders, join in the mixing solutions of the nitric acid of 60 grams, 2 mol concentration and phosphoric acid, wherein the proportioning of nitric acid and phosphoric acid is 1:1, adds the 8 gram vitriol oils again, stirs 2.5 hours at 20-60 ℃, filters, and collects filtrate, and filter residue is standby.Filtrate is adjusted to pH8.0 with sodium hydroxide solution, add the 20% ammoniacal liquor-carbonic acid ammonia solution of 60 grams by equivalent ammoniacal liquor and the preparation of carbonic acid ammonia, in 60 ℃ of stirrings till not producing precipitation, filter, collect filtrate and it is concentrated to 1 mol, electrodeposition promptly obtains 0.381 gram zinc, and the rate of recovery is 95%;
Adding 30% sodium hydroxide solution in above-mentioned filter residue, the weight ratio of filter residue and sodium hydroxide is 0.1-0.5, stirs, and filters, and adds 20 gram sodium sulphite solids again in filtrate, filters to such an extent that black lead sulfide precipitates.Lead sulfide precipitation and solid sodium hydroxide is even with the mixed of 2:3, put into crucible, place resistance furnace to heat then, under 300-400 ℃ temperature, reacted 40-50 minutes, reduce to room temperature, add 1 premium on currency and soaked 20-30 minutes, filter, it is plumbous promptly to get 0.697 gram, and the rate of recovery is 89%.
Embodiment 6
Get the useless mud of 10 grams and be milled down to 250 orders, join in the mixing solutions of the Phenylsulfonic acid of 80 grams, 1.2 mol concentration and hydrochloric acid, wherein the proportioning of Phenylsulfonic acid and hydrochloric acid is 1:1, add 0.8 gram bromine water again, stirred 0.5 hour, filter at 20-60 ℃, collect filtrate, filter residue is standby.Filtrate is adjusted to pH8.0 with sodium hydroxide solution, add the 20% ammoniacal liquor-carbonic acid ammonia solution of 80 grams by equivalent ammoniacal liquor and the preparation of carbonic acid ammonia, in 60 ℃ of stirrings till not producing precipitation, filter, collect filtrate and it is concentrated to 1 mol, electrodeposition promptly obtains 0.369 gram zinc, and the rate of recovery is 93%;
Adding 30% sodium hydroxide solution in above-mentioned filter residue, the weight ratio of filter residue and sodium hydroxide is 0.1-0.5, stirs, and filters, and adds 20 gram sodium sulphite solids again in filtrate, filters to such an extent that black lead sulfide precipitates.Lead sulfide precipitation and solid sodium hydroxide is even with the mixed of 2:3, put into crucible, place resistance furnace to heat then, under 300-400 ℃ temperature, reacted 40-50 minutes, reduce to room temperature, add 1 premium on currency and soaked 20-30 minutes, filter, it is plumbous promptly to get 0.722 gram, and the rate of recovery is 86%.
Embodiment 7
Get the useless mud of 10 grams and be milled down to 250 orders, join in the mixing solutions of the nitric acid of 70 grams, 1.5 mol concentration and tosic acid, wherein the proportioning of nitric acid and tosic acid is 1:1, add 10 gram persulphates again, stirred 4 hours, filter at 20-60 ℃, collect filtrate, filter residue is standby.Filtrate is adjusted to pH8.0 with sodium hydroxide solution, add the 20% ammoniacal liquor-carbonic acid ammonia solution of 70 grams by equivalent ammoniacal liquor and the preparation of carbonic acid ammonia, in 60 ℃ of stirrings till not producing precipitation, filter, collect filtrate and it is concentrated to 1 mol, electrodeposition promptly obtains 0.357 gram zinc, and the rate of recovery is 87%;
Adding 30% sodium hydroxide solution in above-mentioned filter residue, the weight ratio of filter residue and sodium hydroxide is 0.1-0.5, stirs, and filters, and adds 20 gram sodium sulphite solids again in filtrate, filters to such an extent that black lead sulfide precipitates.Lead sulfide precipitation and solid sodium hydroxide is even with the mixed of 2:3, put into crucible, place resistance furnace to heat then, under 300-400 ℃ temperature, reacted 40-50 minutes, reduce to room temperature, add 1 premium on currency and soaked 20-30 minutes, filter, it is plumbous promptly to get 0.706 gram, and the rate of recovery is 84%.
Embodiment 8
Get the useless mud of 10 grams and be milled down to 250 orders, join in the mixing solutions of the phosphoric acid of 40 grams, 1 mol concentration and hydrochloric acid, wherein the proportioning of phosphoric acid and hydrochloric acid is 1:1, adds 1 gram hydrogen peroxide again, stirs 1.5 hours at 20-60 ℃, filters, and collects filtrate, and filter residue is standby.Filtrate is adjusted to pH8.0 with sodium hydroxide solution, add the 20% ammoniacal liquor-carbonic acid ammonia solution of 40 grams by equivalent ammoniacal liquor and the preparation of carbonic acid ammonia, in 60 ℃ of stirrings till not producing precipitation, filter, collect filtrate and it is concentrated to 1 mol, electrodeposition promptly obtains 0.361 gram zinc, and the rate of recovery is 88%;
Adding 30% sodium hydroxide solution in above-mentioned filter residue, the weight ratio of filter residue and sodium hydroxide is 0.1-0.5, stirs, and filters, and adds 20 gram sodium sulphite solids again in filtrate, filters to such an extent that black lead sulfide precipitates.Lead sulfide precipitation and solid sodium hydroxide is even with the mixed of 2:3, put into crucible, place resistance furnace to heat then, under 300-400 ℃ temperature, reacted 40-50 minutes, reduce to room temperature, add 1 premium on currency and soaked 20-30 minutes, filter, it is plumbous promptly to get 0.756 gram, and the rate of recovery is 90%.
Embodiment 9
Get the useless mud of 10 grams and be milled down to 300 orders, join in the mixing solutions of the phosphoric acid of 80 grams, 2 mol concentration and hydrochloric acid, wherein the proportioning of phosphoric acid and hydrochloric acid is 1:1, add 0.5 gram hydrogen peroxide and 0.5 gram persulphate again, stirred 1 hour, filter at 20-60 ℃, collect filtrate, filter residue is standby.Filtrate is adjusted to pH8.0 with sodium hydroxide solution, add the 20% ammoniacal liquor-carbonic acid ammonia solution of 80 grams by equivalent ammoniacal liquor and the preparation of carbonic acid ammonia, in 60 ℃ of stirrings till not producing precipitation, filter, collect filtrate and it is concentrated to 1 mol, electrodeposition promptly obtains 0.353 gram zinc, and the rate of recovery is 86%;
Adding 30% sodium hydroxide solution in above-mentioned filter residue, the weight ratio of filter residue and sodium hydroxide is 0.1-0.5, stirs, and filters, and adds 20 gram sodium sulphite solids again in filtrate, filters to such an extent that black lead sulfide precipitates.Lead sulfide precipitation and solid sodium hydroxide is even with the mixed of 2:3, put into crucible, place resistance furnace to heat then, under 300-400 ℃ temperature, reacted 40-50 minutes, reduce to room temperature, add 1 premium on currency and soaked 20-30 minutes, filter, it is plumbous promptly to get 0.773 gram, and the rate of recovery is 92%.
Claims (3)
1, a kind of method that reclaims zinc lead from waste electrolysis anode sludge is characterized in that comprising the steps:
A. the leaching of waste electrolysis anode sludge
Waste electrolysis anode sludge is ground to the 150-300 order, joins in the mixed acid solution of the doubly useless mud weight of 3-10, the volumetric molar concentration of nitration mixture is the 1-2 mol, adds useless mud weight 0.01-1 oxygenant doubly again, stirs 0.2-4 hour down at 20-60 ℃, filters, and filter residue is standby; Filtrate is adjusted to pH8.0 with sodium hydroxide solution, the ammoniacal liquor-carbonic acid ammonia solution of weight such as adding and nitration mixture, and the weight ratio of ammoniacal liquor and carbonic acid ammonia is 1: 1, the concentration of ammoniacal liquor-carbonic acid ammonia solution is 20%, stirs, and filters, and gets solution of zinc sulfate;
B. electrodeposition
Above-mentioned solution of zinc sulfate is concentrated to 1 mol, after carry out electrodeposition, promptly get metallic zinc;
C. high temperature alkaline purification system lead
In above-mentioned filter residue, add 30% sodium hydroxide solution, the weight ratio of filter residue and sodium hydroxide solution is 0.1-0.5, is stirred to the filtrate clarification, filters, in filtrate, add the sodium sulphite solid again, the weight ratio of sodium sulphite and filter residue is 1:1-2, the lead sulfide black precipitate, lead sulfide precipitation and solid sodium hydroxide are pressed the weight ratio mixing of 2:3, placed 300-400 ℃ of reacting by heating 40-50 minutes, reduce to room temperature, water logging is filtered and is promptly got metallic lead.
2, the method that from waste electrolysis anode sludge, reclaims zinc lead according to claim 1, it is characterized in that: described nitration mixture is the mixture of any two kinds of acid in sulfuric acid, nitric acid, phosphoric acid, Phenylsulfonic acid, hydrochloric acid, the tosic acid, and wherein the weight ratio of two kinds of acid is 1:1.
3, a kind of method that from waste electrolysis anode sludge, reclaims zinc lead according to claim 1, it is characterized in that: described oxygenant is one or both in potassium permanganate, the vitriol oil, perchloric acid, chlorine, industrial bleaching powder, hydrogen peroxide, bromine water, the persulphate, and wherein the weight ratio of two kinds of oxygenants is 1:1.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN200710168561A CN100591783C (en) | 2007-11-29 | 2007-11-29 | Method for recovering zinc and lead from waste electrolytic anode mud |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN200710168561A CN100591783C (en) | 2007-11-29 | 2007-11-29 | Method for recovering zinc and lead from waste electrolytic anode mud |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| CN101451198A true CN101451198A (en) | 2009-06-10 |
| CN100591783C CN100591783C (en) | 2010-02-24 |
Family
ID=40733686
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CN200710168561A Expired - Fee Related CN100591783C (en) | 2007-11-29 | 2007-11-29 | Method for recovering zinc and lead from waste electrolytic anode mud |
Country Status (1)
| Country | Link |
|---|---|
| CN (1) | CN100591783C (en) |
Cited By (16)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN102268699A (en) * | 2011-08-19 | 2011-12-07 | 中南大学 | Method for removing anode slime of porous anode |
| CN102776384A (en) * | 2012-08-08 | 2012-11-14 | 汉源县广超有色金属综合回收有限责任公司 | Method for extracting zinc from non-ferrous metal waste residue |
| CN102912138A (en) * | 2012-10-19 | 2013-02-06 | 湖南有色金属研究院 | Method of recycling zinc, manganese, lead and silver from zinc electrowinning anode mud |
| WO2013111143A1 (en) * | 2012-01-24 | 2013-08-01 | G D Abdhool Rahim | Alloy containing precious metals and its method of preparation by using herbal extracts |
| CN103789554A (en) * | 2014-03-11 | 2014-05-14 | 中南大学 | Copper-nickel slag step-by-step concentration difference leaching method |
| CN104645543A (en) * | 2014-05-04 | 2015-05-27 | 广西梧州鸳江立德粉有限责任公司 | Treatment method of oxidized slag containing zinc and cadmium |
| CN105039691A (en) * | 2015-08-25 | 2015-11-11 | 四川华富宇科技有限责任公司 | Method for leaching low-grade lead-zinc oxide ores through electrooxidation and leaching device of method |
| CN106744964A (en) * | 2016-11-21 | 2017-05-31 | 西北矿冶研究院 | Process for recovering silicon carbide from silicon carbide waste |
| CN106757149A (en) * | 2016-12-28 | 2017-05-31 | 贵州宏达环保科技有限公司 | It is a kind of that manganese, lead, the method for silver are reclaimed from electrolytic zinc anode mud |
| CN107419299A (en) * | 2017-04-17 | 2017-12-01 | 晋宁理工恒达科技有限公司 | It is a kind of to reclaim method of the Zinc electrolysis with the useless plate of palisading type aluminium base alloy lead anode |
| WO2018014748A1 (en) * | 2016-07-19 | 2018-01-25 | 云南祥云飞龙再生科技股份有限公司 | Process for preparing lead with ammonia electroreduction in ammonium chloride |
| CN108624910A (en) * | 2018-04-16 | 2018-10-09 | 北京科技大学 | A kind of zinc Whote-wet method smelting process method of energy-saving and emission-reduction |
| CN109055757A (en) * | 2018-09-10 | 2018-12-21 | 吉首大学 | A method of manganese dioxide and lead in the anode slag of recycling electrolytic manganese or electrolytic zinc |
| CN113604841A (en) * | 2021-07-22 | 2021-11-05 | 田成 | Method for extracting and recovering valuable metal from waste liquid by non-inert anode plate |
| US11319613B2 (en) | 2020-08-18 | 2022-05-03 | Enviro Metals, LLC | Metal refinement |
| CN114561548A (en) * | 2022-03-02 | 2022-05-31 | 金川集团股份有限公司 | Method for removing and recovering lead in high-lead copper anode slime |
-
2007
- 2007-11-29 CN CN200710168561A patent/CN100591783C/en not_active Expired - Fee Related
Cited By (23)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN102268699B (en) * | 2011-08-19 | 2014-11-05 | 中南大学 | Method for removing anode slime of porous anode |
| CN102268699A (en) * | 2011-08-19 | 2011-12-07 | 中南大学 | Method for removing anode slime of porous anode |
| WO2013111143A1 (en) * | 2012-01-24 | 2013-08-01 | G D Abdhool Rahim | Alloy containing precious metals and its method of preparation by using herbal extracts |
| CN102776384A (en) * | 2012-08-08 | 2012-11-14 | 汉源县广超有色金属综合回收有限责任公司 | Method for extracting zinc from non-ferrous metal waste residue |
| CN102912138A (en) * | 2012-10-19 | 2013-02-06 | 湖南有色金属研究院 | Method of recycling zinc, manganese, lead and silver from zinc electrowinning anode mud |
| CN103789554B (en) * | 2014-03-11 | 2015-09-30 | 中南大学 | A kind of substep concentration difference leaching method of copper nickel slag |
| CN103789554A (en) * | 2014-03-11 | 2014-05-14 | 中南大学 | Copper-nickel slag step-by-step concentration difference leaching method |
| CN104645543A (en) * | 2014-05-04 | 2015-05-27 | 广西梧州鸳江立德粉有限责任公司 | Treatment method of oxidized slag containing zinc and cadmium |
| CN104645543B (en) * | 2014-05-04 | 2017-11-17 | 广西梧州鸳江立德粉有限责任公司 | Processing method containing zinc, the oxidizing slag of cadmium |
| CN105039691A (en) * | 2015-08-25 | 2015-11-11 | 四川华富宇科技有限责任公司 | Method for leaching low-grade lead-zinc oxide ores through electrooxidation and leaching device of method |
| WO2018014748A1 (en) * | 2016-07-19 | 2018-01-25 | 云南祥云飞龙再生科技股份有限公司 | Process for preparing lead with ammonia electroreduction in ammonium chloride |
| US10563315B2 (en) | 2016-07-19 | 2020-02-18 | Xiangyun Tenglong Investment Co., Ltd. | Process for preparing lead by electroreduction with ammonium chloride and ammonia |
| CN106744964A (en) * | 2016-11-21 | 2017-05-31 | 西北矿冶研究院 | Process for recovering silicon carbide from silicon carbide waste |
| CN106757149A (en) * | 2016-12-28 | 2017-05-31 | 贵州宏达环保科技有限公司 | It is a kind of that manganese, lead, the method for silver are reclaimed from electrolytic zinc anode mud |
| CN107419299B (en) * | 2017-04-17 | 2019-06-07 | 晋宁理工恒达科技有限公司 | A method of the recycling Zinc electrolysis useless plate of palisading type aluminium base alloy lead anode |
| CN107419299A (en) * | 2017-04-17 | 2017-12-01 | 晋宁理工恒达科技有限公司 | It is a kind of to reclaim method of the Zinc electrolysis with the useless plate of palisading type aluminium base alloy lead anode |
| CN108624910A (en) * | 2018-04-16 | 2018-10-09 | 北京科技大学 | A kind of zinc Whote-wet method smelting process method of energy-saving and emission-reduction |
| CN109055757A (en) * | 2018-09-10 | 2018-12-21 | 吉首大学 | A method of manganese dioxide and lead in the anode slag of recycling electrolytic manganese or electrolytic zinc |
| CN109055757B (en) * | 2018-09-10 | 2020-04-03 | 吉首大学 | Method for recovering manganese dioxide and lead in anode slag of electrolytic manganese or electrolytic zinc |
| US11319613B2 (en) | 2020-08-18 | 2022-05-03 | Enviro Metals, LLC | Metal refinement |
| US11578386B2 (en) | 2020-08-18 | 2023-02-14 | Enviro Metals, LLC | Metal refinement |
| CN113604841A (en) * | 2021-07-22 | 2021-11-05 | 田成 | Method for extracting and recovering valuable metal from waste liquid by non-inert anode plate |
| CN114561548A (en) * | 2022-03-02 | 2022-05-31 | 金川集团股份有限公司 | Method for removing and recovering lead in high-lead copper anode slime |
Also Published As
| Publication number | Publication date |
|---|---|
| CN100591783C (en) | 2010-02-24 |
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| CN100591783C (en) | Method for recovering zinc and lead from waste electrolytic anode mud | |
| CN102491287B (en) | Process for separating and recovering selenium from selenium-containing material | |
| CN101565174B (en) | Method for extracting refined tellurium from tellurium-contained smelting slag | |
| CN101289171B (en) | Hydrometallurgy process for abstracting 99.99% Te and 99.99% Bi2O3 and comprehensively recovering accessory products | |
| CN101503761B (en) | Method for separating and recycling valuable metal from pressure leached high sulphur slag | |
| CN108624759B (en) | Method for comprehensively recovering valuable metals from white smoke | |
| CN102382980B (en) | Method for directly purifying cadmium from sponge cadmium | |
| CN113684368A (en) | Method for co-processing arsenic sulfide slag and arsenic-containing smoke dust in copper smelting | |
| CN105463197A (en) | Method for recycling valuable metal from copper smelting white smoke | |
| CN105905874A (en) | Method of recycling tellurium from anode mud | |
| US9175411B2 (en) | Gold and silver extraction technology | |
| CN106636656A (en) | Method for recovering valuable metal from copper smelting white smoke | |
| CN114318417B (en) | Method for producing electrolytic manganese by low-grade manganese oxide three-ore method | |
| CN101328539A (en) | Oxidation oven ash hydrometallurgical leaching process | |
| CN103757421A (en) | Method for extracting rare and precious metals from platinum-palladium mud | |
| CN112458280A (en) | Method for extracting valuable metals by leaching low grade nickel matte with acidic etching solution | |
| CN102690947A (en) | Smelting process of silver concentrate | |
| CN103290212A (en) | Clean smelting method of indium | |
| CN102757022B (en) | Technology for extracting tellurium product and valuable metal from lead anode slime | |
| CN1239720C (en) | Method for producing high purity metal zinc from zinc oxide ore | |
| CN105967153A (en) | Technology for recovering tellurium from high-tellurium slag | |
| CN106542506A (en) | A kind of method that selenium is reclaimed from heavy tellurium waste liquid | |
| CN104005051A (en) | Method for extracting electrodeposited zinc from zinc oxide powder in ammonium chloride solution system | |
| CN102002597B (en) | Method for comprehensively recovering valuable metals from low-grade tellurium slag | |
| CN113481371A (en) | Method for efficiently recovering antimony, bismuth, copper and silver from silver separating slag of lead anode slime |
Legal Events
| Date | Code | Title | Description |
|---|---|---|---|
| C06 | Publication | ||
| PB01 | Publication | ||
| C10 | Entry into substantive examination | ||
| SE01 | Entry into force of request for substantive examination | ||
| C14 | Grant of patent or utility model | ||
| GR01 | Patent grant | ||
| C17 | Cessation of patent right | ||
| CF01 | Termination of patent right due to non-payment of annual fee |
Granted publication date: 20100224 Termination date: 20101129 |