CN1119678A - Method for recovering silver, gold, antimony, copper and lead from lead anode mud - Google Patents
Method for recovering silver, gold, antimony, copper and lead from lead anode mud Download PDFInfo
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- CN1119678A CN1119678A CN94104747A CN94104747A CN1119678A CN 1119678 A CN1119678 A CN 1119678A CN 94104747 A CN94104747 A CN 94104747A CN 94104747 A CN94104747 A CN 94104747A CN 1119678 A CN1119678 A CN 1119678A
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- leaching
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- 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
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Abstract
A method for recovering silver, gold, antimony, copper and lead from lead anode slime uses hydrochloric acid to leach out antimony and copper, uses sodium chlorate-hydrochloric acid-sodium chloride to leach out gold and uses ammonia to leach out spongy silver. Its leaching time is short, leaching effect is good, acid consumption is low and cost is low, and its direct-yield rate is up to above 96% and silver content is 99.3%. Said process can directly obtain analytically pure silver nitrate.
Description
The invention relates to a method for recovering silver, gold, antimony, copper and lead from lead anode slime, belonging to the technical field of hydrometallurgy.
The recovery of associated silver from anode slime of nonferrous metals such as lead, copper, nickel and the like has become an important source of silver, and China always adopts thetraditional pyrometallurgical treatment of the anode slime for many years. The pyrogenic process has the characteristics of high processing capacity, strong raw material adaptability and the like. But also has the defects of long process flow, much return slag, large investment, long production period and the like. And a large amount of flue gas such as arsenic, lead, antimony, selenium and the like is discharged in the process of reducing smelting and oxidizing refining by a pyrogenic process, so that the environment is seriously polluted. In recent years, a plurality of scientific research institutes and smelting plants in China research out the method for treating the anode mud by the wet method, the process flow is simpler, the number of intermediate products is less, the investment for building the plant is less, the capital turnover is fast, and the useful metals in the anode mud can be effectively and comprehensively utilized. For example, chinese patent 90109944.9 discloses a method for preparing silver nitrate and recovering copper, lead and antimony from lead anode slime, which comprises leaching the filter residue after nitric acid leaching with hydrochloric acid, precipitating lead with sulfuric acid, hydrolyzing hydrochloric acid leaching solution, washing the filter residue obtained by hydrolysis with alkali, precipitating silver with hydrochloric acid, ammonia leaching, complexing, reducing, etc. It also has the defects of complex working procedure, large consumption of nitric acid, high cost, low recovery rate of silver and the like.
The invention aims to provide a method for recovering silver, gold, antimony, copper and lead from various lead anode slime, which has the advantages of low acid consumption, low cost, high silver recovery rate and low pollution and can overcome the defects in the prior art.
The purpose of the invention is realized by the following technicalsolution: the method comprises a preparation process for leaching copper and antimony by hydrochloric acid, and the specific process comprises the following steps:
1. leaching for 1-2 hours at 60-90 ℃ by using hydrochloric acid with the liquid-solid ratio of 3: 1-6: 1 and the 3N-5.5N, removing copper and antimony in the anode mud, recovering antimony through hydrolysis, and replacing copper with scrap iron;
2. adding sodium chlorate accounting for 3-10% of the weight of the residue into the filter residue, leaching the filter residue for 1-3 hours at 60-90 ℃ by using hydrochloric acid accounting for 0.3-1N and sodium chloride accounting for 2-12% of the weight of the residue in a liquid-solid ratio of 3: 1-5: 1, filtering, and reducing the filtrate by using oxalic acid or ferric sulfite or sulfur dioxide to obtain sponge gold;
3. leaching the sodium chlorate leaching residue for 1-3 hours by using ammonia water with the liquid-solid ratio of 3: 1-6: 1 and the liquid-solid ratio of 3N-5N at 20-60 ℃, filtering, and adding excessive activated carbon or activated aluminum with the amount of 1.2-2 times into the filtrate to reduce to obtain the sponge silver.
The attached drawing is a process flow chart of the invention.
The technical solution of the present invention will now be further described with reference to the accompanying drawings.
First, hydrochloric acid soaks antimony, copper
The antimony in the lead anode slime is mainly in the form of metallic antimony and antimony trichloride, and only a small amount of antimony exists in the form of antimony alloy. Adding hydrochloric acid with the liquid-solid ratio of 5: 1 and the concentration of 4N into the lead anode slime, and leaching for 1-1.5 hours at the temperature of 75-90 ℃. In the leaching process, antimony trioxide reacts with hydrochloric acid to form an antimony trichloride solution such as
Metallic antimony also reacts with hydrochloric acid to form an antimony trichloride solution, e.g.
The copper in the anode slime is mainly present as metallic copper and partly as basic copper carbonate, the copper and basic copper carbonate react with hydrochloric acid to form cuprous chloride solution, e.g. copper chloride
In the process, more than 95 percent of antimony and copper are leached, and antimony and copper chloride are combined and separated with silver-containing filtrate by vacuum filtration, wherein the leaching rate of silver is lower than 1 percent.
Chlorination gold leaching
After antimony and copper are leached by hydrochloric acid, the leaching residue mainly contains silver, gold and lead compounds, 4-6% of excessive sodium chlorate calculated by the weight of the residue is added into the hydrochloric acid medium, 6-10% of sodium chloride calculated by the weight of the residue and hydrochloric acid with the concentration of 0.5-0.6N and the liquid-solid ratio of 4: 1 are added to leach gold, the leaching temperature is 75-85 ℃, and the leaching time is 1.5-2.5 hours. The following reactions take place in the process:
the action of sodium chlorate and hydrochloric acid produces hypochlorous acid, which is immediately decomposed to produce active oxygen to oxidize gold and silver. At the moment, gold forms complex ions to enter the solution, and oxidized silver is combined with oxygen and chlorine to generate silver chloride precipitate with low solubility product to enter the slag. The concentration of industrial sodium chloride is about 40g/L and the concentration of hydrochloric acid medium is 0.5N-0.6N when gold leaching is carried out by chlorination, so that the [ AgCl]is difficult to form2]-So silver is hardly leached out. During chlorination, the copper content in the slag is further reduced. The purpose of chlorination is to convert substantially all of the silver to silver chloride.
Three, ammonia leaching process
After chlorination, most of silver in the slag is converted into silver chloride, lead is converted into lead dichloride, ammonia water with the concentration of 3-4N and the liquid-solid ratio of 5: 1 is added into the slag for secondary ammonia leaching, the leaching temperature is 40-55 ℃, and the time is 1.2-1.8 hours. And (3) carrying out secondary ammonia leaching on the filter residue after the primary ammonia leaching to ensure that the silver chloride is completely converted into ammonia-silver complex ions, wherein the reaction is as follows:
since the lead dichloride is not complexed with the ammonia water, the ammonia leaching solution is Ag (NH)3)2 +。
Four, silver reduction process
Ag (NH) produced after complexing reaction of silver chloride3)2Reducing in a stainless steel reaction kettle by using dechlorinating agent such as activated carbon or activated aluminum with the excess of 1.5 times, wherein the reduction process comprises the following steps:
filtering the reduced sponge silver, washing the sponge silver by secondary distilled water, and drying the sponge silver to obtain the sponge silver with the silver content of more than 99.0 percent.
The sponge silver is washed and then reacts with reagent-grade silver nitrate to generate silver nitrate, and the reaction formula is as follows:
adding 11-13 liters of 65% reagent-grade silver nitrate into 20 liters of secondary distilled water in a reaction kettle, reacting with 8-11 kilograms of sponge silver, and evaporating, crystallizing, filtering and drying to obtain the reagent-grade silver nitrate.
The results of the pilot test are shown in the following table:
| material weighing apparatus | Weight of material Kg | Filtrate L | Filtering residue Kg | Slag grade% | The content of the solution is g/l | The leaching rate is% | Recovery rate | |||||
| Cu | Sb | Ag | Cu | Sb | Ag | Cu | Sb | Ag | ||||
| Anode mud | 100 | 6.1 | 9.50 | 16.40 | ||||||||
| Copper-leaching antimony-leaching slag | 43.68 | 0.04 | 0.34 | 37.60 | ||||||||
| Copper leaching and antimony leaching solution | 490 | 12.41 | 19.08 | Not detected out | 99.68 | 98.41 | ||||||
| Chlorination slag | 40 | 0.04 | 0.31 | |||||||||
| Chlorination solution | Not detected out | |||||||||||
| Ammonia leaching residue | 25.31 | Pb 62.10 | 0.1 | |||||||||
| Ammonia leaching solution | 470 | 33.5 | ||||||||||
| Sponge silver | 15.9 | 99.30 | 96.27 | |||||||||
| Reducing liquid | 460 | 0.1 | ||||||||||
The invention overcomes the defects in the background technology, improves the leaching effect of copper and antimony by using hydrochloric acid as a leaching agent, does not need to add an oxidant, and can effectively separate the copper and the antimony from silver, gold and lead. In the process of leaching gold from sodium chlorate-hydrochloric acid-sodium chloride, hypochlorous acid generated after the action of sodium chlorate and hydrochloric acid is immediately decomposed into active oxygen, gold and silver are oxidized, gold forms complex ions and enters into solution, and silver, oxygen and chlorine are combined to generate silver chloride precipitate with low solubility product, and the silver chloride precipitate exists in slag and is ready for complexing silver by ammonia. The silver-ammonia complex generated after complexing the silver chloride is reduced to sponge silver by using activated carbon or activated aluminum with the excess of 1.5 times, and the direct recovery rate of the silver is more than 96 percent. The silver nitrate can be directly prepared into analytically pure or industrial pure silver nitrate by adding silver nitrate after the sponge silver is washed. The method has the advantages of short leaching time, good leaching effect, low acid consumption, low cost, less pollution, capability of producing the silver nitrate of analytical pure grade meeting the GB 670-86 standard, and capability of producing the silver nitrate of the sponge silver with the silver content of more than 99.3 percent, and is suitable for treating various lead anode slime.
Claims (4)
1. A method for recovering silver, gold, antimony, copper and lead from lead anode slime comprises the step of leaching antimony by using hydrochloric acid to leach copper, and is characterized in that:
1.1, leaching for 1-2 hours at 60-90 ℃ by using hydrochloric acid with the liquid-solid ratio of 3: 1-6: 1 and the 3N-5.5N, removing copper and antimonyin anode mud, recovering antimony through hydrolysis, and replacing copper with scrap iron;
1.2 adding sodium chlorate accounting for 3-10% of the weight of the residue into the filter residue, leaching the filter residue for 1-3 hours at 60-90 ℃ by using hydrochloric acid accounting for 0.3-1N and sodium chloride accounting for 2-12% of the weight of the residue in a liquid-solid ratio of 3: 1-5: 1, filtering, and reducing the filtrate into sponge gold by using oxalic acid, ferric sulfite or sulfur dioxide;
1.3, leaching the sodium chlorate leaching residue for 1-3 hours by using ammonia water with the liquid-solid ratio of 3: 1-6: 1 and the liquid-solid ratio of 3N-5N at 20-60 ℃, filtering, and adding excessive activated carbon or activated aluminum with the amount of 1.2-2 times into the filtrate to reduce to obtain the sponge silver.
2. The method as claimed in claim 1, wherein the liquid-solid ratio of the hydrochloric acid leaching in 1.1 is 4: 1-6: 1, the concentration is 4N-5N, and the leaching temperature is preferably 75-90 ℃.
3. The method as set forth in claim 1, wherein the weight of said sodium chlorate in 1.2 is 5-10% of the weight of the residue, the concentration of hydrochloric acid is 0.5-0.8N, the liquid-solid ratio is 3.5: 1-4.5: 1, the amount of sodium chloride is 4-11% of the weight of the residue, the leaching temperature is 70-90 ℃, and the leaching time is 1-2.5 hours.
4. The method according to claim 1, wherein the ammonia water concentration of 1.3 is 3 to 4N, the liquid-solid ratio is 4: 1 to 5: 1, and the ammonia leaching is carried out at 30 to 55 ℃.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN94104747A CN1119678A (en) | 1994-05-10 | 1994-05-10 | Method for recovering silver, gold, antimony, copper and lead from lead anode mud |
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| Application Number | Priority Date | Filing Date | Title |
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| CN94104747A CN1119678A (en) | 1994-05-10 | 1994-05-10 | Method for recovering silver, gold, antimony, copper and lead from lead anode mud |
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| CN1119678A true CN1119678A (en) | 1996-04-03 |
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| CN94104747A Pending CN1119678A (en) | 1994-05-10 | 1994-05-10 | Method for recovering silver, gold, antimony, copper and lead from lead anode mud |
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Cited By (10)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN1045794C (en) * | 1996-09-16 | 1999-10-20 | 昆明贵金属研究所 | Method for obtg. gold, silver and other noble metals from anode and contg. high concn. arsenic and lead |
| CN1051114C (en) * | 1996-10-25 | 2000-04-05 | 昆明贵金属研究所 | Pretreatment method before wet method acid immersion for lead anode mud |
| EP1061143A1 (en) * | 1999-06-18 | 2000-12-20 | W.C. Heraeus GmbH & Co. KG | Method for solubilizing noble metalls |
| CN101787440A (en) * | 2010-03-29 | 2010-07-28 | 广西华锡集团股份有限公司 | Method for wet processing of high antimony-lead anode mud |
| CN102031383A (en) * | 2010-11-28 | 2011-04-27 | 郴州市金贵银业股份有限公司 | Wet process for lead-silver residues |
| CN102079545A (en) * | 2010-06-02 | 2011-06-01 | 李德良 | Preparation method of ferric chloride |
| CN102925703A (en) * | 2012-11-22 | 2013-02-13 | 湖南稀土金属材料研究院 | Method for recycling valuable metals from lead anode slime |
| CN109112304A (en) * | 2017-06-22 | 2019-01-01 | 生命地带有限公司 | It is a kind of to extract expensive, low-priced, rare element hydrometallurgical process method |
| CN109593969A (en) * | 2018-12-14 | 2019-04-09 | 许金刚 | A kind of molten gold process of no cyanogen |
| CN110205500A (en) * | 2019-06-18 | 2019-09-06 | 山东黄金冶炼有限公司 | The removal methods of impurity in a kind of reduction bronze |
-
1994
- 1994-05-10 CN CN94104747A patent/CN1119678A/en active Pending
Cited By (15)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN1045794C (en) * | 1996-09-16 | 1999-10-20 | 昆明贵金属研究所 | Method for obtg. gold, silver and other noble metals from anode and contg. high concn. arsenic and lead |
| CN1051114C (en) * | 1996-10-25 | 2000-04-05 | 昆明贵金属研究所 | Pretreatment method before wet method acid immersion for lead anode mud |
| EP1061143A1 (en) * | 1999-06-18 | 2000-12-20 | W.C. Heraeus GmbH & Co. KG | Method for solubilizing noble metalls |
| US6440376B1 (en) | 1999-06-18 | 2002-08-27 | W. C. Heraeus Gmbh & Co. Kg. | Method for dissolving noble metals out of segregated ores containing noble metals |
| CN101787440A (en) * | 2010-03-29 | 2010-07-28 | 广西华锡集团股份有限公司 | Method for wet processing of high antimony-lead anode mud |
| CN101787440B (en) * | 2010-03-29 | 2011-08-10 | 广西华锡集团股份有限公司 | Method for wet processing of high antimony-lead anode mud |
| CN102079545A (en) * | 2010-06-02 | 2011-06-01 | 李德良 | Preparation method of ferric chloride |
| CN102079545B (en) * | 2010-06-02 | 2013-01-09 | 李德良 | Preparation method of ferric chloride |
| CN102031383A (en) * | 2010-11-28 | 2011-04-27 | 郴州市金贵银业股份有限公司 | Wet process for lead-silver residues |
| CN102031383B (en) * | 2010-11-28 | 2012-07-25 | 郴州市金贵银业股份有限公司 | Wet process for lead-silver residues |
| CN102925703A (en) * | 2012-11-22 | 2013-02-13 | 湖南稀土金属材料研究院 | Method for recycling valuable metals from lead anode slime |
| CN109112304A (en) * | 2017-06-22 | 2019-01-01 | 生命地带有限公司 | It is a kind of to extract expensive, low-priced, rare element hydrometallurgical process method |
| CN109593969A (en) * | 2018-12-14 | 2019-04-09 | 许金刚 | A kind of molten gold process of no cyanogen |
| CN110205500A (en) * | 2019-06-18 | 2019-09-06 | 山东黄金冶炼有限公司 | The removal methods of impurity in a kind of reduction bronze |
| CN110205500B (en) * | 2019-06-18 | 2021-09-17 | 山东黄金冶炼有限公司 | Method for removing impurities in reduced gold powder |
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