CN103088218A - Method for extracting silver and lead from smelting slag generated by pyrogenic process treatment of copper anode mud - Google Patents
Method for extracting silver and lead from smelting slag generated by pyrogenic process treatment of copper anode mud Download PDFInfo
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- 239000002893 slag Substances 0.000 title claims abstract description 63
- 229910052709 silver Inorganic materials 0.000 title claims abstract description 57
- 239000004332 silver Substances 0.000 title claims abstract description 57
- 238000003723 Smelting Methods 0.000 title claims abstract description 50
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 title claims abstract description 42
- 229910052802 copper Inorganic materials 0.000 title claims abstract description 41
- 239000010949 copper Substances 0.000 title claims abstract description 41
- 238000000034 method Methods 0.000 title claims abstract description 30
- 230000001698 pyrogenic effect Effects 0.000 title 1
- 238000002386 leaching Methods 0.000 claims abstract description 102
- 238000000605 extraction Methods 0.000 claims abstract description 16
- AKHNMLFCWUSKQB-UHFFFAOYSA-L sodium thiosulfate Chemical compound [Na+].[Na+].[O-]S([O-])(=O)=S AKHNMLFCWUSKQB-UHFFFAOYSA-L 0.000 claims abstract description 13
- 235000019345 sodium thiosulphate Nutrition 0.000 claims abstract description 13
- GEHJYWRUCIMESM-UHFFFAOYSA-L sodium sulfite Chemical compound [Na+].[Na+].[O-]S([O-])=O GEHJYWRUCIMESM-UHFFFAOYSA-L 0.000 claims abstract description 12
- JVBXVOWTABLYPX-UHFFFAOYSA-L sodium dithionite Chemical compound [Na+].[Na+].[O-]S(=O)S([O-])=O JVBXVOWTABLYPX-UHFFFAOYSA-L 0.000 claims abstract description 9
- VHUUQVKOLVNVRT-UHFFFAOYSA-N Ammonium hydroxide Chemical compound [NH4+].[OH-] VHUUQVKOLVNVRT-UHFFFAOYSA-N 0.000 claims abstract description 7
- 235000011114 ammonium hydroxide Nutrition 0.000 claims abstract description 7
- 229910000365 copper sulfate Inorganic materials 0.000 claims abstract description 7
- ARUVKPQLZAKDPS-UHFFFAOYSA-L copper(II) sulfate Chemical compound [Cu+2].[O-][S+2]([O-])([O-])[O-] ARUVKPQLZAKDPS-UHFFFAOYSA-L 0.000 claims abstract description 7
- 235000010265 sodium sulphite Nutrition 0.000 claims abstract description 6
- SRWFBFUYENBCGF-UHFFFAOYSA-M sodium;chloride;hydrochloride Chemical compound [Na+].Cl.[Cl-] SRWFBFUYENBCGF-UHFFFAOYSA-M 0.000 claims abstract description 6
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 claims description 54
- 238000000926 separation method Methods 0.000 claims description 50
- 239000007788 liquid Substances 0.000 claims description 48
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 claims description 33
- FAPWRFPIFSIZLT-UHFFFAOYSA-M Sodium chloride Chemical compound [Na+].[Cl-] FAPWRFPIFSIZLT-UHFFFAOYSA-M 0.000 claims description 28
- 239000000243 solution Substances 0.000 claims description 23
- 238000003756 stirring Methods 0.000 claims description 15
- 239000000203 mixture Substances 0.000 claims description 14
- 239000011780 sodium chloride Substances 0.000 claims description 14
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims description 11
- 239000000126 substance Substances 0.000 claims description 10
- 239000010802 sludge Substances 0.000 claims description 9
- HWSZZLVAJGOAAY-UHFFFAOYSA-L lead(II) chloride Chemical compound Cl[Pb]Cl HWSZZLVAJGOAAY-UHFFFAOYSA-L 0.000 claims description 6
- -1 sodium thiosulfate-copper sulfate-sodium sulfite Chemical compound 0.000 claims description 5
- 239000007787 solid Substances 0.000 claims description 5
- 239000011259 mixed solution Substances 0.000 claims description 4
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 claims 1
- 239000011593 sulfur Substances 0.000 claims 1
- 229910052717 sulfur Inorganic materials 0.000 claims 1
- 239000000706 filtrate Substances 0.000 abstract description 6
- 239000000284 extract Substances 0.000 abstract description 3
- 238000009776 industrial production Methods 0.000 abstract description 2
- 238000005272 metallurgy Methods 0.000 abstract description 2
- 229910052751 metal Inorganic materials 0.000 description 12
- 239000002184 metal Substances 0.000 description 11
- 150000002739 metals Chemical class 0.000 description 7
- PCHJSUWPFVWCPO-UHFFFAOYSA-N gold Chemical compound [Au] PCHJSUWPFVWCPO-UHFFFAOYSA-N 0.000 description 6
- 229910052737 gold Inorganic materials 0.000 description 6
- 239000010931 gold Substances 0.000 description 6
- 239000010970 precious metal Substances 0.000 description 6
- DGAQECJNVWCQMB-PUAWFVPOSA-M Ilexoside XXIX Chemical compound C[C@@H]1CC[C@@]2(CC[C@@]3(C(=CC[C@H]4[C@]3(CC[C@@H]5[C@@]4(CC[C@@H](C5(C)C)OS(=O)(=O)[O-])C)C)[C@@H]2[C@]1(C)O)C)C(=O)O[C@H]6[C@@H]([C@H]([C@@H]([C@H](O6)CO)O)O)O.[Na+] DGAQECJNVWCQMB-PUAWFVPOSA-M 0.000 description 4
- 239000000047 product Substances 0.000 description 4
- 238000004064 recycling Methods 0.000 description 4
- 239000011734 sodium Substances 0.000 description 4
- 229910052708 sodium Inorganic materials 0.000 description 4
- 238000007654 immersion Methods 0.000 description 3
- 239000000843 powder Substances 0.000 description 3
- 229910052714 tellurium Inorganic materials 0.000 description 3
- PORWMNRCUJJQNO-UHFFFAOYSA-N tellurium atom Chemical compound [Te] PORWMNRCUJJQNO-UHFFFAOYSA-N 0.000 description 3
- JPVYNHNXODAKFH-UHFFFAOYSA-N Cu2+ Chemical compound [Cu+2] JPVYNHNXODAKFH-UHFFFAOYSA-N 0.000 description 2
- PXHVJJICTQNCMI-UHFFFAOYSA-N Nickel Chemical compound [Ni] PXHVJJICTQNCMI-UHFFFAOYSA-N 0.000 description 2
- KDLHZDBZIXYQEI-UHFFFAOYSA-N Palladium Chemical compound [Pd] KDLHZDBZIXYQEI-UHFFFAOYSA-N 0.000 description 2
- BUGBHKTXTAQXES-UHFFFAOYSA-N Selenium Chemical compound [Se] BUGBHKTXTAQXES-UHFFFAOYSA-N 0.000 description 2
- 230000009286 beneficial effect Effects 0.000 description 2
- 229910001431 copper ion Inorganic materials 0.000 description 2
- 230000000694 effects Effects 0.000 description 2
- 238000003912 environmental pollution Methods 0.000 description 2
- 150000002500 ions Chemical class 0.000 description 2
- BASFCYQUMIYNBI-UHFFFAOYSA-N platinum Chemical compound [Pt] BASFCYQUMIYNBI-UHFFFAOYSA-N 0.000 description 2
- 239000002994 raw material Substances 0.000 description 2
- 229910052711 selenium Inorganic materials 0.000 description 2
- 239000011669 selenium Substances 0.000 description 2
- 150000004760 silicates Chemical class 0.000 description 2
- 239000002910 solid waste Substances 0.000 description 2
- 150000003467 sulfuric acid derivatives Chemical class 0.000 description 2
- DJHGAFSJWGLOIV-UHFFFAOYSA-K Arsenate3- Chemical class [O-][As]([O-])([O-])=O DJHGAFSJWGLOIV-UHFFFAOYSA-K 0.000 description 1
- VEXZGXHMUGYJMC-UHFFFAOYSA-M Chloride anion Chemical compound [Cl-] VEXZGXHMUGYJMC-UHFFFAOYSA-M 0.000 description 1
- ZAMOUSCENKQFHK-UHFFFAOYSA-N Chlorine atom Chemical compound [Cl] ZAMOUSCENKQFHK-UHFFFAOYSA-N 0.000 description 1
- XUIMIQQOPSSXEZ-UHFFFAOYSA-N Silicon Chemical compound [Si] XUIMIQQOPSSXEZ-UHFFFAOYSA-N 0.000 description 1
- 230000002411 adverse Effects 0.000 description 1
- 239000000956 alloy Substances 0.000 description 1
- 229910045601 alloy Inorganic materials 0.000 description 1
- 229910052787 antimony Inorganic materials 0.000 description 1
- WATWJIUSRGPENY-UHFFFAOYSA-N antimony atom Chemical compound [Sb] WATWJIUSRGPENY-UHFFFAOYSA-N 0.000 description 1
- 229910052788 barium Inorganic materials 0.000 description 1
- DSAJWYNOEDNPEQ-UHFFFAOYSA-N barium atom Chemical compound [Ba] DSAJWYNOEDNPEQ-UHFFFAOYSA-N 0.000 description 1
- 229910052797 bismuth Inorganic materials 0.000 description 1
- JCXGWMGPZLAOME-UHFFFAOYSA-N bismuth atom Chemical compound [Bi] JCXGWMGPZLAOME-UHFFFAOYSA-N 0.000 description 1
- 238000005266 casting Methods 0.000 description 1
- 238000005660 chlorination reaction Methods 0.000 description 1
- 239000000460 chlorine Substances 0.000 description 1
- 229910052801 chlorine Inorganic materials 0.000 description 1
- 239000012141 concentrate Substances 0.000 description 1
- 238000005868 electrolysis reaction Methods 0.000 description 1
- 239000003792 electrolyte Substances 0.000 description 1
- 238000005265 energy consumption Methods 0.000 description 1
- 230000007613 environmental effect Effects 0.000 description 1
- 230000003631 expected effect Effects 0.000 description 1
- 239000012535 impurity Substances 0.000 description 1
- 239000011133 lead Substances 0.000 description 1
- 150000002611 lead compounds Chemical class 0.000 description 1
- 238000004519 manufacturing process Methods 0.000 description 1
- 238000010299 mechanically pulverizing process Methods 0.000 description 1
- 238000002844 melting Methods 0.000 description 1
- 230000008018 melting Effects 0.000 description 1
- 229910052759 nickel Inorganic materials 0.000 description 1
- 239000007800 oxidant agent Substances 0.000 description 1
- 229910052763 palladium Inorganic materials 0.000 description 1
- 229910052697 platinum Inorganic materials 0.000 description 1
- 238000011084 recovery Methods 0.000 description 1
- 238000007670 refining Methods 0.000 description 1
- 150000003839 salts Chemical class 0.000 description 1
- 229910052710 silicon Inorganic materials 0.000 description 1
- 239000010703 silicon Substances 0.000 description 1
- 229940100890 silver compound Drugs 0.000 description 1
- 150000003379 silver compounds Chemical class 0.000 description 1
- OGFYIDCVDSATDC-UHFFFAOYSA-N silver silver Chemical compound [Ag].[Ag] OGFYIDCVDSATDC-UHFFFAOYSA-N 0.000 description 1
- 150000004763 sulfides Chemical class 0.000 description 1
- 239000002699 waste material Substances 0.000 description 1
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
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- Manufacture And Refinement Of Metals (AREA)
Abstract
本发明提供了一种从铜阳极泥火法处理产生的熔炼渣中提取银、铅的方法，属于冶金技术领域。该方法是将铜阳极泥火法处理产生的熔炼渣进行破碎，先采用硫代硫酸钠、硫酸铜、亚硫酸钠和氨水作为浸出液浸出提取银；浸出滤液采用连二亚硫酸钠还原后可返回继续浸出银；对浸出银后的浸出渣采用盐酸-氯化钠浸出提取铅，浸出铅滤液析出分离铅后返回继续浸出铅。本发明工艺简单，操作方便，工艺可靠，银、铅的提取率高（银的浸出率为75~87%，铅的提取率为90%以上）最终滤液循环使用，成本低，适应性强，环境友好，适合大规模工业生产，具有很好的应用前景。The invention provides a method for extracting silver and lead from smelting slag produced by mud fire treatment of copper anodes, belonging to the technical field of metallurgy. The method is to crush the smelting slag produced by the copper anode slime treatment, first use sodium thiosulfate, copper sulfate, sodium sulfite and ammonia water as the leaching solution to leach and extract silver; the leaching filtrate can be returned to continue leaching silver after being reduced by sodium dithionite; Use hydrochloric acid-sodium chloride to extract lead from the leaching residue after leaching silver, and return to continue leaching lead after leaching lead filtrate to precipitate and separate lead. The invention has the advantages of simple process, convenient operation, reliable process, high extraction rate of silver and lead (the extraction rate of silver is 75-87%, and the extraction rate of lead is more than 90%). The final filtrate is recycled, with low cost and strong adaptability. It is environmentally friendly, suitable for large-scale industrial production, and has good application prospects.
Description
技术领域 technical field
本发明属于冶金技术领域，涉及铜阳极泥火法处理过程中产生的固体废渣的回收利用，尤其涉及一种从铜阳极泥火法处理产生的熔炼渣中提取银、铅的方法。 The invention belongs to the technical field of metallurgy, and relates to the recycling of solid waste slag produced during copper anode slime treatment, in particular to a method for extracting silver and lead from the smelting slag produced by copper anode slime treatment.
背景技术 Background technique
铜阳极泥是在对铜精矿的电解精炼过程中，由比铜电位更正的元素和不溶于电解液的各种物质组成，其成分主要取决于铜阳极的组成、铸造质量以及电解的技术条件等。分银渣是铜阳极泥在提取贵金属金、银、铂、钯和铜、硒、碲等有价元素后的余渣，而熔炼渣是铜阳极泥在经加压强化浸出镍、铜、碲后再干燥、火法熔炼得到金银粗合金过程中产生的残渣。熔炼渣中主要含有钠、硅、钡、铅等元素，另还含有一定的稀贵金属元素，如金、银、硒、碲、铋、铂族金属等，也是提取贵金属的重要原料。熔炼渣中化学成分及相组成均非常复杂，大多为复杂的硅酸盐、砷酸盐、硫酸盐、氧化物和钠的金属酸盐，其中铅含量在10~20%（质量百分比），银的质量百分含量在0.1~1.0%。铜阳极泥火法处理产生的熔炼渣是综合回收银、金、铅等有价金属的良好原料，对其进行回收再利用，对于解决贵金属资源的严重匮乏有重要的意义。 Copper anode slime is composed of elements more positive than copper potential and various substances insoluble in electrolyte during the electrolytic refining process of copper concentrate. Its composition mainly depends on the composition of copper anode, casting quality and electrolysis technical conditions, etc. . Silver separation slag is the residual slag of copper anode slime after extracting precious metals such as gold, silver, platinum, palladium and copper, selenium, tellurium and other valuable elements, while smelting slag is the copper anode slime after the leaching of nickel, copper, tellurium and so on. Then dry and pyromelt the residue produced in the process of obtaining gold and silver crude alloy. Smelting slag mainly contains sodium, silicon, barium, lead and other elements, and also contains certain rare and precious metal elements, such as gold, silver, selenium, tellurium, bismuth, platinum group metals, etc., which are also important raw materials for extracting precious metals. The chemical composition and phase composition of the smelting slag are very complex, mostly complex metal salts of silicates, arsenates, sulfates, oxides and sodium, of which the lead content is 10-20% (mass percentage), silver The mass percentage content of 0.1~1.0%. The smelting slag produced by copper anode slime treatment is a good raw material for the comprehensive recovery of valuable metals such as silver, gold, and lead. Its recycling and reuse is of great significance to solve the serious shortage of precious metal resources.
由于熔炼渣成分复杂，较之脉石矿，对其进行独立的贵金属提取与回收，选矿难度更大，若将其直接返回选矿厂选矿则达不到预期效果；而较之阳极泥，渣中贵金属含量低，处理更加困难。目前，国内外对熔炼渣的处理方式为将其直接返回熔炼炉以提取金、银，不但金银收率较低，还会造成铅、锑等危害及杂质元素浸出，从而影响后续阴极铜的质量。另外，只对熔炼渣中的某种单一金属元素的进行火法还原熔炼提取，由于熔炼渣中含有高熔点的硅酸盐和硫酸盐，火法还原熔炼提取单一金属元素不但带来高能耗、低效率的不良影响，从而既浪费了渣中的某些有价金属，又造成了新的环境污染。采用湿法对熔炼渣进行综合回收利用，由于铜阳极泥火法处理产生的熔炼渣成分相当复杂，综合回收利用熔炼渣中的大多数有价金属周期长、工艺复杂繁重、效率低。 Due to the complex composition of smelting slag, compared with gangue ore, it is more difficult to extract and recover precious metals independently, and if it is directly returned to the beneficiation plant for beneficiation, the expected effect will not be achieved; The lower precious metal content makes handling more difficult. At present, domestic and foreign smelting slags are treated by directly returning them to the smelting furnace to extract gold and silver. Not only is the yield of gold and silver low, but it will also cause hazards such as lead and antimony and the leaching of impurity elements, thereby affecting the subsequent production of cathode copper. quality. In addition, only a certain single metal element in the smelting slag is extracted by pyro-reduction smelting. Since the smelting slag contains silicates and sulfates with high melting points, the extraction of a single metal element by pyro-reduction smelting not only brings high energy consumption, The adverse effects of low efficiency not only waste some valuable metals in the slag, but also cause new environmental pollution. The wet method is used to comprehensively recycle smelting slag. Since the composition of smelting slag produced by copper anode slime treatment is quite complex, the comprehensive recycling of most of the valuable metals in smelting slag takes a long period, complicated and heavy process, and low efficiency.
发明内容 Contents of the invention
本发明的目的是针对现有技术中存在的问题，提供一种从铜阳极泥火法处理产生的熔炼渣中提取银、铅的方法。 The purpose of the present invention is to provide a method for extracting silver and lead from the smelting slag produced by copper anode sludge fire treatment in order to solve the problems in the prior art.
本发明从铜阳极泥火法处理产生的熔炼渣中提取银、铅的方法，包括以下工艺步骤： The present invention extracts the method for silver, lead from the smelting slag that copper anode slime treatment produces, comprises following process step:
（1）浸出提银：将铜阳极泥火法处理产生的熔炼渣进行破碎后，采用硫代硫酸钠-硫酸铜-亚硫酸钠溶液的混合液作为浸出液，控制固液比在1:6 ~1:10g/ml；用氨水调pH=8~11，在20~45℃的温度下搅拌浸出0.5~5h，固液分离得到浸出滤渣和浸出液；浸出滤渣用于浸出提铅；浸出液中加入连二亚硫酸钠，在50~70℃下搅拌还原0.5~2h，固液分离，分离渣为粗银粉，分离液中添加硫代硫酸钠（加入量为0.01~0.1 g/ml）调整后返回继续浸出银。 (1) Silver extraction by leaching: After crushing the smelting slag produced by the mud fire treatment of copper anodes, a mixture of sodium thiosulfate-copper sulfate-sodium sulfite solution is used as the leaching solution, and the solid-liquid ratio is controlled at 1:6 ~1: 10g/ml; use ammonia water to adjust pH=8~11, stir and leaching at 20~45℃ for 0.5~5h, separate solid and liquid to obtain leaching filter residue and leaching solution; leaching filter residue is used for leaching and extracting lead; add sodium dithionite to the leaching solution , stirring and reducing at 50~70°C for 0.5~2h, solid-liquid separation, the separation residue is coarse silver powder, adding sodium thiosulfate (addition amount: 0.01~0.1 g/ml) to the separation liquid to adjust and return to continue leaching silver.
所述混合浸出液中，硫代硫酸钠可与银离子形成较稳定的络合离子，对银的化合物具有良好的浸出效果，且浸出速率快，选择性浸出效率高；硫酸铜中的铜离子具有氧化剂的作用，氨水既可以调节pH值，又可以与铜离子形成络合离子，有利于银的浸出。 In the mixed leaching solution, sodium thiosulfate can form relatively stable complex ions with silver ions, and has a good leaching effect on silver compounds, and the leaching rate is fast, and the selective leaching efficiency is high; the copper ions in copper sulfate have As an oxidizing agent, ammonia water can not only adjust the pH value, but also form complex ions with copper ions, which is beneficial to the leaching of silver.
所述混合浸出液中，硫代硫酸钠的质量为熔炼渣质量的20~50%；硫酸铜的质量为熔炼渣质量的2~6%，亚硫酸钠的质量为熔炼渣质量的4~12%；所述连二亚硫酸钠的加入量为熔炼渣质量的1~5%。银的浸出率为75~87%。 In the mixed leachate, the quality of sodium thiosulfate is 20% to 50% of the quality of smelting slag; the quality of copper sulfate is 2% to 6% of the quality of smelting slag, and the quality of sodium sulfite is 4% to 12% of the quality of smelting slag; The amount of sodium dithionite added is 1-5% of the mass of smelting slag. The leaching rate of silver is 75~87%.
（2）浸出提铅：将步骤（1）的浸出滤渣采用浓盐酸-氯化钠的混合液作为浸出液，控制固液比在1:6 ~1:10g/ml，在90~100℃下搅拌浸出1~6h，固液分离得到浸出渣和浸出液；浸出渣经检验化学成分后可继续用于提取其他金属等它用；浸出液加水稀释1~10倍析出富铅物，再进行固液分离得分离渣和分离液；分离渣为高含量的氯化铅，可用于生产铅产品；分离液中添加盐酸和氯化钠调整后（盐酸的添加量为0.05~0.2 g/ml，氯化钠的添加量为0.1~0.3 g/ml）后返回继续浸出铅。 (2) Lead extraction by leaching: use the mixture of concentrated hydrochloric acid-sodium chloride as the leaching solution for the leaching filter residue of step (1), control the solid-liquid ratio at 1:6~1:10g/ml, and stir at 90~100°C After leaching for 1-6 hours, solid-liquid separation is carried out to obtain leaching slag and leaching liquid; the leaching slag can be used to extract other metals and other purposes after the chemical composition of the leaching slag is checked; Separation slag and separation liquid; separation slag is lead chloride with high content, which can be used to produce lead products; hydrochloric acid and sodium chloride are added to the separation liquid after adjustment (the addition amount of hydrochloric acid is 0.05~0.2 g/ml, the amount of sodium chloride The addition amount is 0.1~0.3 g/ml) and return to continue leaching lead.
在浓盐酸-氯化钠的混合液中，浓盐酸可将铜阳极泥火法分银后的残渣中的氧化物、硫化物等转化成可溶性的盐酸盐，有利于铅的浸出；氯化钠主要是提高溶液中的氯离子浓度，又可将单质铅及铅的化合物转化为铅的氯化物，用于铅的浸出。 In the mixed solution of concentrated hydrochloric acid-sodium chloride, concentrated hydrochloric acid can convert the oxides and sulfides in the residue after copper anode slime silver separation into soluble hydrochloride, which is beneficial to the leaching of lead; chlorination Sodium is mainly used to increase the concentration of chloride ions in the solution, and it can also convert elemental lead and lead compounds into lead chloride for the leaching of lead.
在浓盐酸-氯化钠的混合液中，所述浓盐酸的终浓度为50~200g/L，氯化钠的终浓度为100~300g/L。铅的提取率高达95%以上。 In the concentrated hydrochloric acid-sodium chloride mixed solution, the final concentration of the concentrated hydrochloric acid is 50-200g/L, and the final concentration of sodium chloride is 100-300g/L. The extraction rate of lead is as high as 95%.
本发明相对现有技术具有以下优点： The present invention has the following advantages relative to the prior art:
1、本发明将铜阳极泥火法处理产生的熔炼渣进行破碎后，先采用硫代硫酸钠-硫酸铜-亚硫酸钠-氨水作为浸出液浸出提取银，浸出滤液采用连二亚硫酸钠还原银后可返回继续浸出银；对浸出银后的浸出渣采用盐酸-氯化钠浸出提取铅，浸出铅滤液析出分离铅后返回继续浸出铅，银和铅的提取率高，实现了铜阳极泥火法处理过程中产生的固体废渣的回收利用。 1. In the present invention, after crushing the smelting slag produced by copper anode sludge fire treatment, sodium thiosulfate-copper sulfate-sodium sulfite-ammonia water is used as the leaching solution to leach and extract silver, and the leaching filtrate can be returned to continue after using sodium dithionite to reduce silver Silver leaching; use hydrochloric acid-sodium chloride leaching to extract lead from the leaching residue after leaching silver, leach lead filtrate to separate lead and return to continue leaching lead, the extraction rate of silver and lead is high, and the process of copper anode slime fire treatment Recycling of generated solid waste.
2、本发明工艺简单，操作方便，最终滤液循环使用，节约了成本，减少了环境污染。 2. The process of the present invention is simple, the operation is convenient, and the final filtrate is recycled, which saves costs and reduces environmental pollution.
3、本发明设备简单，工艺可靠，环境友好，适应性强，适合大规模工业生产，具有很好的应用前景。 3. The invention has the advantages of simple equipment, reliable process, environmental friendliness, strong adaptability, suitable for large-scale industrial production, and has good application prospects.
具体实施方式 Detailed ways
下面通过具体实施例对本发明从铜阳极泥火法处理产生的熔炼渣中提取银、铅的方法作详细说明。 The method for extracting silver and lead in the present invention from the smelting slag produced by the mud fire treatment of copper anodes will be described in detail below through specific examples.
实施例1 Example 1
（1）浸出提银：将铜阳极泥火法处理产生的熔炼渣进行机械粉碎后，准确称量残渣粉末100g，放入1000ml烧杯中，加入600ml水（固液比为1:6 g/ml），添加4g硫酸铜、10g亚硫酸钠、50g硫代硫酸钠和20ml氨水（系统pH =10.8），加热到35℃，并保持体系在33~37℃，持续搅拌浸银2h；然后进行液固分离，得到浸出滤渣（91.8g）和浸出液；浸出滤渣进入下一步骤进行浸出提铅；浸出液中加入3g连二亚硫酸钠，升温到60℃，并保持体系在55~65℃，持续搅拌还原银30min，固液分离，分离渣为粗银粉；分离液中可再添加20g硫代硫酸钠继续返回进行浸银。铜阳极泥火法处理产生的熔炼渣中银的浸出率为87.0%。 (1) Silver extraction by leaching: After mechanically pulverizing the smelting slag produced by the copper anode sludge fire treatment, accurately weigh 100g of the residue powder, put it into a 1000ml beaker, add 600ml of water (solid-liquid ratio is 1:6 g/ml ), add 4g copper sulfate, 10g sodium sulfite, 50g sodium thiosulfate and 20ml ammonia water (system pH = 10.8), heat to 35°C, keep the system at 33~37°C, keep stirring and immerse silver for 2h; then carry out liquid-solid separation , to obtain leaching filter residue (91.8g) and leaching solution; the leaching filter residue enters the next step for leaching and extracting lead; add 3g of sodium dithionite to the leaching solution, raise the temperature to 60°C, and keep the system at 55~65°C, and keep stirring to reduce silver for 30min. Solid-liquid separation, the separation residue is coarse silver powder; 20g of sodium thiosulfate can be added to the separation liquid to continue to return for silver immersion. The leaching rate of silver in the smelting slag produced by copper anode slime treatment was 87.0%.
（2）浸出提铅：将上述提银浸出滤渣放入1000ml烧杯中，加入700ml水（固液比为1:8 g/ml），分别加入浓盐酸、氯化钠，搅拌，浓盐酸的终浓度为80g/L，氯化钠浓度为250g/L；升温到95℃，并保持体系在90℃以上，搅拌浸出2h；过滤固液分离得到浸出渣和浸出液；浸出渣经检验化学成分后可继续用于提取其他金属等它用；浸出液中加水稀释2倍析出富铅物，再进行固液分离得分离渣和分离液，分离渣为高含量的氯化铅，可用于生产铅产品；分离后液添加80g盐酸和200g氯化钠后返回继续浸出铅。铜阳极泥火法处理产生的熔炼渣中铅的浸出率为95.3%。 (2) Lead extraction by leaching: put the above-mentioned silver leaching filter residue into a 1000ml beaker, add 700ml of water (solid-to-liquid ratio is 1:8 g/ml), add concentrated hydrochloric acid and sodium chloride respectively, stir, and the final concentration of concentrated hydrochloric acid The concentration is 80g/L, and the concentration of sodium chloride is 250g/L; the temperature is raised to 95°C, and the system is kept above 90°C, and stirred and leached for 2 hours; the solid-liquid separation is filtered to obtain leaching residue and leaching liquid; the leaching residue can be tested for chemical composition Continue to be used to extract other metals, etc.; add water to the leachate to dilute 2 times to precipitate lead-rich substances, and then conduct solid-liquid separation to obtain separation slag and separation liquid. The separation slag is high-content lead chloride, which can be used to produce lead products; separation Add 80g of hydrochloric acid and 200g of sodium chloride to the back liquid and return to continue leaching lead. The leaching rate of lead in the smelting slag produced by the slime treatment of copper anode was 95.3%.
实施例2 Example 2
（1）浸出提银：将铜阳极泥火法处理产生的熔炼渣进行机械粉碎，准确称量残渣粉末100g，放入1000ml烧杯中，加入800ml水（固液比为1:8 g/ml），添加4g硫酸铜、8g亚硫酸钠、40g硫代硫酸钠和15ml氨水（系统pH =10.3），加热到35℃，并保持体系在33~37℃，持续搅拌浸银2h。然后进行液固分离得到浸出滤渣（92.1g）和浸出液；浸出滤渣进入下一步骤进行浸出提铅；浸出液中加入3g连二亚硫酸钠，升温到60℃，并保持体系在55~65℃，持续搅拌还原银30min，固液分离，分离渣为粗银粉；分离液中可再添加25g硫代硫酸钠继续返回进行浸银。铜阳极泥火法处理产生的熔炼渣中银的浸出率为81.9%。 (1) Silver extraction by leaching: mechanically pulverize the smelting slag produced by copper anode mud fire treatment, accurately weigh 100g of residue powder, put it into a 1000ml beaker, add 800ml of water (solid-liquid ratio is 1:8 g/ml) , add 4g of copper sulfate, 8g of sodium sulfite, 40g of sodium thiosulfate and 15ml of ammonia water (system pH = 10.3), heat to 35°C, keep the system at 33~37°C, and keep stirring to immerse in silver for 2h. Then carry out liquid-solid separation to obtain leaching filter residue (92.1g) and leaching solution; leaching filter residue enters the next step for leaching and extracting lead; add 3g of sodium dithionite to the leaching solution, raise the temperature to 60°C, and keep the system at 55~65°C, and keep stirring Silver reduction for 30 minutes, solid-liquid separation, and the separation residue is coarse silver powder; 25g of sodium thiosulfate can be added to the separation liquid to continue to return to silver immersion. The leaching rate of silver in the smelting slag produced by the slime treatment of copper anode was 81.9%.
（2）浸出提铅：将上述提银浸出滤渣放入1000ml烧杯中，加入700ml水（固液比为1:8 g/ml），分别加入浓盐酸、氯化钠，搅拌，浓盐酸的终浓度为64g/L，氯化钠浓度为250g/L，升温到95℃，并保持体系在90℃以上，搅拌浸出2h，过滤固液分离得到浸出渣和浸出液；经检验化学成分后可继续用于提取其他金属等它用；浸出液加水稀释2倍析出富铅物，再进行固液分离得分离渣和分离液，分离渣为高含量的氯化铅，可用于生产铅产品；分离液添加100g盐酸和250g氯化钠后返回继续浸出铅。将铜阳极泥火法处理产生的熔炼渣中铅的浸出率为98.3%。 (2) Lead extraction by leaching: put the above-mentioned silver leaching filter residue into a 1000ml beaker, add 700ml of water (solid-to-liquid ratio is 1:8 g/ml), add concentrated hydrochloric acid and sodium chloride respectively, stir, and the final concentration of concentrated hydrochloric acid The concentration is 64g/L, the concentration of sodium chloride is 250g/L, the temperature is raised to 95°C, and the system is kept above 90°C, stirred and leached for 2 hours, and the solid-liquid separation is filtered to obtain the leaching residue and leaching liquid; after checking the chemical composition, it can continue to be used It is used for extracting other metals, etc.; dilute the leaching solution with water twice to precipitate lead-rich substances, and then conduct solid-liquid separation to obtain separation slag and separation liquid. The separation slag contains high content of lead chloride and can be used to produce lead products; add 100g to the separation liquid After hydrochloric acid and 250g sodium chloride return to continue leaching lead. The leaching rate of lead in the smelting slag produced by mud fire treatment of copper anode was 98.3%.
实施例3 Example 3
（1）浸出提银：将铜阳极泥火法处理产生的熔炼渣进行机械粉碎，准确称量残渣粉末500g，放入5000ml烧杯中，加入3000ml水（固液比为1:6 g/ml），添加20g硫酸铜、45g亚硫酸钠、200g硫代硫酸钠和70ml氨水（系统pH =10.5），加热到35℃，并保持体系在32~39℃，持续搅拌浸银1h。然后进行液固分离得到浸出滤渣（461.8 g）和浸出液；浸出滤渣进入下一步骤进行浸出提铅；浸出液中加入15g连二亚硫酸钠，升温到60℃，并保持体系在52~68℃，持续搅拌还原银45min，固液分离，分离渣为粗银粉，分离液中可再添加125g硫代硫酸钠继续返回进行浸银。铜阳极泥火法处理产生的熔炼渣中银的浸出率为83.7%。 (1) Silver extraction by leaching: mechanically pulverize the smelting slag produced by copper anode mud fire treatment, accurately weigh 500g of residue powder, put it into a 5000ml beaker, add 3000ml of water (solid-liquid ratio is 1:6 g/ml) , add 20g of copper sulfate, 45g of sodium sulfite, 200g of sodium thiosulfate and 70ml of ammonia water (system pH = 10.5), heat to 35°C, keep the system at 32~39°C, and keep stirring to immerse in silver for 1h. Then carry out liquid-solid separation to obtain leaching filter residue (461.8 g) and leaching solution; leaching filter residue enters the next step for leaching and extracting lead; add 15 g of sodium dithionite to the leaching solution, raise the temperature to 60°C, and keep the system at 52~68°C, and keep stirring Silver reduction for 45 minutes, solid-liquid separation, the separation residue is coarse silver powder, and 125g sodium thiosulfate can be added to the separation liquid to continue to return to silver immersion. The leaching rate of silver in the smelting slag produced by the slime treatment of copper anode was 83.7%.
（2）浸出提铅：将上述提银浸出滤渣放入5000ml烧杯中，加入3600ml水（固液比为1:8 g/ml），分别加入浓盐酸、氯化钠，搅拌，浓盐酸的终浓度为80g/L，氯化钠浓度为250g/L，升温到95℃，并保持体系在90℃以上，搅拌浸出2h，过滤固液分离得到浸出渣和浸出液；浸出渣经检验化学成分后可继续用于提取其他金属等它用；浸出液加水稀释2倍析出富铅物，再进行固液分离，分离渣为高含量的氯化铅，可用于生产铅产品；分离液添加300g盐酸和650g氯化钠后返回继续浸出铅。铜阳极泥火法处理产生的熔炼渣中铅的浸出率为98.6%。 (2) Lead extraction by leaching: put the above-mentioned silver leaching filter residue into a 5000ml beaker, add 3600ml of water (solid-to-liquid ratio is 1:8 g/ml), add concentrated hydrochloric acid and sodium chloride respectively, stir, and the final concentration of concentrated hydrochloric acid The concentration is 80g/L, the concentration of sodium chloride is 250g/L, the temperature is raised to 95°C, and the system is kept above 90°C, stirred and leached for 2 hours, and the solid-liquid separation is filtered to obtain the leaching residue and leaching liquid; the leaching residue can be tested for chemical composition. Continue to be used to extract other metals, etc.; dilute the leaching solution with water twice to precipitate lead-rich substances, and then perform solid-liquid separation. The separation residue is high-content lead chloride, which can be used to produce lead products; add 300g hydrochloric acid and 650g chlorine to the separation solution Return to continue leaching lead after sodium is dissolved. The leaching rate of lead in the smelting slag produced by the slime treatment of copper anode was 98.6%.
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