CN1058299C - Ag and Au extracting and Sb, Bi, Cu and Pb recovering process from lead slime - Google Patents
Ag and Au extracting and Sb, Bi, Cu and Pb recovering process from lead slime Download PDFInfo
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- 229910052709 silver Inorganic materials 0.000 title claims abstract description 40
- 229910052787 antimony Inorganic materials 0.000 title claims abstract description 23
- 229910052797 bismuth Inorganic materials 0.000 title claims abstract description 20
- 229910052737 gold Inorganic materials 0.000 title claims abstract description 19
- 238000000034 method Methods 0.000 title claims abstract description 18
- 229910052802 copper Inorganic materials 0.000 title claims abstract description 15
- 229910052745 lead Inorganic materials 0.000 title abstract 2
- 239000004332 silver Substances 0.000 claims abstract description 33
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 claims abstract description 32
- 239000010949 copper Substances 0.000 claims abstract description 24
- 238000002386 leaching Methods 0.000 claims abstract description 24
- PCHJSUWPFVWCPO-UHFFFAOYSA-N gold Chemical compound [Au] PCHJSUWPFVWCPO-UHFFFAOYSA-N 0.000 claims abstract description 20
- WATWJIUSRGPENY-UHFFFAOYSA-N antimony atom Chemical compound [Sb] WATWJIUSRGPENY-UHFFFAOYSA-N 0.000 claims abstract description 19
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims abstract description 16
- JCXGWMGPZLAOME-UHFFFAOYSA-N bismuth atom Chemical compound [Bi] JCXGWMGPZLAOME-UHFFFAOYSA-N 0.000 claims abstract description 16
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 claims abstract description 15
- 229910021607 Silver chloride Inorganic materials 0.000 claims abstract description 7
- HKZLPVFGJNLROG-UHFFFAOYSA-M silver monochloride Chemical compound [Cl-].[Ag+] HKZLPVFGJNLROG-UHFFFAOYSA-M 0.000 claims abstract description 7
- 238000003723 Smelting Methods 0.000 claims abstract description 6
- 239000002253 acid Substances 0.000 claims abstract description 4
- 239000010931 gold Substances 0.000 claims description 29
- 239000007788 liquid Substances 0.000 claims description 27
- 239000002893 slag Substances 0.000 claims description 24
- 238000003756 stirring Methods 0.000 claims description 20
- FAPWRFPIFSIZLT-UHFFFAOYSA-M Sodium chloride Chemical compound [Na+].[Cl-] FAPWRFPIFSIZLT-UHFFFAOYSA-M 0.000 claims description 16
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 15
- CDBYLPFSWZWCQE-UHFFFAOYSA-L Sodium Carbonate Chemical compound [Na+].[Na+].[O-]C([O-])=O CDBYLPFSWZWCQE-UHFFFAOYSA-L 0.000 claims description 12
- VWDWKYIASSYTQR-UHFFFAOYSA-N sodium nitrate Chemical compound [Na+].[O-][N+]([O-])=O VWDWKYIASSYTQR-UHFFFAOYSA-N 0.000 claims description 12
- 238000000926 separation method Methods 0.000 claims description 11
- XEEYBQQBJWHFJM-UHFFFAOYSA-N iron Substances [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims description 10
- 239000012452 mother liquor Substances 0.000 claims description 9
- 239000007787 solid Substances 0.000 claims description 9
- MUBZPKHOEPUJKR-UHFFFAOYSA-N Oxalic acid Chemical compound OC(=O)C(O)=O MUBZPKHOEPUJKR-UHFFFAOYSA-N 0.000 claims description 8
- 239000011780 sodium chloride Substances 0.000 claims description 8
- 229910000029 sodium carbonate Inorganic materials 0.000 claims description 6
- 235000010344 sodium nitrate Nutrition 0.000 claims description 6
- 239000004317 sodium nitrate Substances 0.000 claims description 5
- 238000010438 heat treatment Methods 0.000 claims description 4
- QZPSXPBJTPJTSZ-UHFFFAOYSA-N aqua regia Chemical compound Cl.O[N+]([O-])=O QZPSXPBJTPJTSZ-UHFFFAOYSA-N 0.000 claims description 3
- WABPQHHGFIMREM-UHFFFAOYSA-N lead(0) Chemical compound [Pb] WABPQHHGFIMREM-UHFFFAOYSA-N 0.000 claims description 3
- 239000000463 material Substances 0.000 claims description 3
- 239000002244 precipitate Substances 0.000 claims description 3
- 238000007670 refining Methods 0.000 claims description 3
- 238000006243 chemical reaction Methods 0.000 claims description 2
- 229910001385 heavy metal Inorganic materials 0.000 claims description 2
- 150000002500 ions Chemical class 0.000 claims description 2
- HWSZZLVAJGOAAY-UHFFFAOYSA-L lead(II) chloride Chemical compound Cl[Pb]Cl HWSZZLVAJGOAAY-UHFFFAOYSA-L 0.000 claims description 2
- 230000007935 neutral effect Effects 0.000 claims description 2
- PDWVXNLUDMQFCH-UHFFFAOYSA-N oxoantimony;hydrochloride Chemical compound Cl.[Sb]=O PDWVXNLUDMQFCH-UHFFFAOYSA-N 0.000 claims description 2
- BWOROQSFKKODDR-UHFFFAOYSA-N oxobismuth;hydrochloride Chemical compound Cl.[Bi]=O BWOROQSFKKODDR-UHFFFAOYSA-N 0.000 claims description 2
- 239000008399 tap water Substances 0.000 claims description 2
- 235000020679 tap water Nutrition 0.000 claims description 2
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims description 2
- 238000007654 immersion Methods 0.000 claims 2
- 235000006408 oxalic acid Nutrition 0.000 claims 2
- 239000000956 alloy Substances 0.000 claims 1
- 229910045601 alloy Inorganic materials 0.000 claims 1
- 238000009835 boiling Methods 0.000 claims 1
- AXCZMVOFGPJBDE-UHFFFAOYSA-L calcium dihydroxide Chemical compound [OH-].[OH-].[Ca+2] AXCZMVOFGPJBDE-UHFFFAOYSA-L 0.000 claims 1
- 239000000920 calcium hydroxide Substances 0.000 claims 1
- 229910001861 calcium hydroxide Inorganic materials 0.000 claims 1
- 235000011116 calcium hydroxide Nutrition 0.000 claims 1
- JJLJMEJHUUYSSY-UHFFFAOYSA-L copper(II) hydroxide Inorganic materials [OH-].[OH-].[Cu+2] JJLJMEJHUUYSSY-UHFFFAOYSA-L 0.000 claims 1
- AEJIMXVJZFYIHN-UHFFFAOYSA-N copper;dihydrate Chemical compound O.O.[Cu] AEJIMXVJZFYIHN-UHFFFAOYSA-N 0.000 claims 1
- 239000000706 filtrate Substances 0.000 claims 1
- 238000001914 filtration Methods 0.000 claims 1
- 238000001556 precipitation Methods 0.000 abstract description 13
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 abstract description 9
- 235000011149 sulphuric acid Nutrition 0.000 abstract description 5
- 238000005868 electrolysis reaction Methods 0.000 abstract description 4
- 230000007062 hydrolysis Effects 0.000 abstract description 4
- 238000006460 hydrolysis reaction Methods 0.000 abstract description 4
- 238000006386 neutralization reaction Methods 0.000 abstract description 2
- 150000003839 salts Chemical class 0.000 abstract description 2
- 238000006073 displacement reaction Methods 0.000 abstract 2
- 238000004090 dissolution Methods 0.000 abstract 1
- 239000001117 sulphuric acid Substances 0.000 abstract 1
- 238000011084 recovery Methods 0.000 description 12
- 238000002791 soaking Methods 0.000 description 4
- 235000017550 sodium carbonate Nutrition 0.000 description 4
- 229910001316 Ag alloy Inorganic materials 0.000 description 2
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 description 2
- 239000006227 byproduct Substances 0.000 description 2
- 238000004519 manufacturing process Methods 0.000 description 2
- 239000000843 powder Substances 0.000 description 2
- 235000008733 Citrus aurantifolia Nutrition 0.000 description 1
- 235000011941 Tilia x europaea Nutrition 0.000 description 1
- 229910052785 arsenic Inorganic materials 0.000 description 1
- RQNWIZPPADIBDY-UHFFFAOYSA-N arsenic atom Chemical compound [As] RQNWIZPPADIBDY-UHFFFAOYSA-N 0.000 description 1
- 239000010953 base metal Substances 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 229910052681 coesite Inorganic materials 0.000 description 1
- 229910052906 cristobalite Inorganic materials 0.000 description 1
- 230000008021 deposition Effects 0.000 description 1
- 238000005516 engineering process Methods 0.000 description 1
- 230000007613 environmental effect Effects 0.000 description 1
- 238000009854 hydrometallurgy Methods 0.000 description 1
- 239000004615 ingredient Substances 0.000 description 1
- 239000013067 intermediate product Substances 0.000 description 1
- 229910052742 iron Inorganic materials 0.000 description 1
- 239000004571 lime Substances 0.000 description 1
- 238000005272 metallurgy Methods 0.000 description 1
- 239000000203 mixture Substances 0.000 description 1
- 238000007254 oxidation reaction Methods 0.000 description 1
- 239000010970 precious metal Substances 0.000 description 1
- 239000002994 raw material Substances 0.000 description 1
- 230000008929 regeneration Effects 0.000 description 1
- 238000011069 regeneration method Methods 0.000 description 1
- 229910052710 silicon Inorganic materials 0.000 description 1
- 239000010703 silicon Substances 0.000 description 1
- 239000000377 silicon dioxide Substances 0.000 description 1
- 235000012239 silicon dioxide Nutrition 0.000 description 1
- 150000003378 silver Chemical class 0.000 description 1
- OGFYIDCVDSATDC-UHFFFAOYSA-N silver silver Chemical compound [Ag].[Ag] OGFYIDCVDSATDC-UHFFFAOYSA-N 0.000 description 1
- 229910052682 stishovite Inorganic materials 0.000 description 1
- 238000006467 substitution reaction Methods 0.000 description 1
- 229910052905 tridymite Inorganic materials 0.000 description 1
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Abstract
Description
从铅阳极泥提取金、银及回收锑、铋、铜、铅的方法。本发明是关于贵金属冶金。The invention discloses a method for extracting gold and silver from lead anode slime and recovering antimony, bismuth, copper and lead. This invention relates to precious metal metallurgy.
铅阳极泥是提取金、银，特别是银的主要原料之一，我国银产量的二分之一以上来自铅阳极泥。铅阳极泥主要成分变动范围(％)：Au 0.002-0.8，Ag 0.1-25，Cu 0.5-10，Bi 1-20 ，As 0.5-27，Sb 0.1-43，S 0.1-2.13，Te 0.1-0.74。铅阳极泥传统处理方法为火法工艺，经还原熔炼-氧化吹炼-金银合金电解得银-银阳极泥回收金。火法的优点是对物料适应性强，生产规模大，设备简单。缺点是一部分金、银在中间产物积压，因此金、银直收率低，金近93％，银85-88％，同时锑、铋、铜、铅也分散，回收率低。有一工厂铅阳极泥成份为(％)：Au 0.10，Ag 10.09，Pb 15.96，Sb 32.69，Bi 1.99，Cu 8.85，As 0.36，Fe 0.16，Ca 0.70，Mg 0.034，SiO2 22.48，属于低砷高硅铅阳极泥，贵贱金属比约为1∶5.8。原来采用火法处理。Lead anode slime is one of the main raw materials for extracting gold and silver, especially silver. More than half of my country's silver production comes from lead anode slime. Change range of main components of lead anode slime (%): Au 0.002-0.8, Ag 0.1-25, Cu 0.5-10, Bi 1-20, As 0.5-27, Sb 0.1-43, S 0.1-2.13, Te 0.1-0.74 . The traditional treatment method of lead anode slime is pyrotechnics, through reduction smelting-oxidation blowing-gold-silver alloy electrolysis to obtain silver-silver anode slime and recover gold. The advantages of the fire method are strong adaptability to materials, large production scale and simple equipment. The disadvantage is that some gold and silver are backlogged in the intermediate products, so the direct recovery rate of gold and silver is low, nearly 93% for gold and 85-88% for silver. At the same time, antimony, bismuth, copper, and lead are also dispersed, and the recovery rate is low. The composition of lead anode slime in a factory is (%): Au 0.10, Ag 10.09, Pb 15.96, Sb 32.69, Bi 1.99, Cu 8.85, As 0.36, Fe 0.16, Ca 0.70, Mg 0.034, SiO2 22.48, belonging to low arsenic and high silicon lead For anode slime, the ratio of noble and base metals is about 1:5.8. Originally treated with fire.
本发明的目的，是针对上述工厂铅阳极泥，研究探索出湿法处理铅阳级泥，得到富银渣，金、银直收率高(Au 99％，Ag 98％)，同时可以综合回收锑、铋、铜、铅，其回收率锑、铋＞80％，铜)65％的方法。本发明所提供的方法，包括盐酸+硫酸混酸浸出(或氯化钠+硫酸介质浸出)，水解沉锑、水解沉铋、中和置换沉铜，盐浸脱铅，置换氯化银、沉铅、熔炼银电解得银，王水溶解金还原金几个步骤，其方法步骤是：1、将铅阳极泥在HCl 2-4mol/L-+H2SO4 2mol/L(或NaCl 3.5-4.5mol/L+H2SO4 2-4mol/L)介质，温度80-90℃，物料重量克数与液体体积毫升数之比为1∶6-10条件，搅拌浸出2-4h，锑、铋、铜转入溶液，固液分离；2将1步骤所得浸出液在温度为常温，搅拌加入自来水至体积稀释到2倍，并加入少量10-20％(重量比)NaOH溶液调整溶液PH值为1-1.5条件水解锑，至SbOCl沉淀完毕，固液分离；3、将2步骤所得沉锑母液，在温度为常温，搅拌条件下，加入10-20％(重量比)NaOH溶液调整溶液PH值到3.5-4.0水解铋，至BiOCl沉淀完全，固液分离；4、将3步骤所得沉铋母液，在温度为常温，搅拌条件下加入10-20％(重量比)NaOH溶液，调整溶液PH到6.5-7.0，至Cu(OH)2沉淀完全；或在温度为常温，搅拌条件下，按Cu∶Fe＝1∶1-1.2(重量比)，向溶液中加入铁粉置换铜，反应1-2h，固液分离；5，将1步骤所得浸出渣，在PH 2-4，温度为近80℃，渣重量克数与溶液体积毫升数之比为1∶13-18条件，用浓度为150-200g/L的NaCl溶液，搅拌浸出2 h，使铅完全转化为氯化铅浸出，固液分离；6、将5步骤所得浸铅液，在温度60-70℃，搅拌条件下，按每1m3溶液加入铅粉2-4kg(或铁粉4-6kg)置换溶于浸铅液中氯化银，反应1-2h，固液分离；接着在温度为常温，搅拌条件下，用石灰[Ca(OH)2或CaO]中和置换液至PH为9-11，使铅及重金属离子和硫酸根沉淀，固液分离，沉铅母液用盐酸回调PH至2-4，即可返回5步骤起始应用；7、将5步骤所得浸铅渣与6步骤所得置换银渣合并，在温度1050-1100℃，加入碳酸钠及硝酸钠进行熔炼，渣∶碳酸钠∶硝酸钠＝1∶1.2-2.5∶0.05-0.1(重量比)，时间10-60min，熔炼得到含金粗银锭，粗银用传统电解精炼法得到纯银；8、将7步骤所得到银电解精炼阳极泥，加入王水溶解，银阳极泥与王水的量比(重量比)为1∶3-4。分次加入，自热或后期加热、搅动，使金溶解完全，银以氯化银形式留于渣中，过滤分离，滤液赶硝、赶酸后，在沸腾温度，加入固体草酸还原金，草酸加入的量比为H2C2O4∶Au＝1-3∶1，时间4-6h，海绵金粉用3mol/L HCl煮洗，再用水洗至中性，得到纯金。The purpose of the present invention is to research and explore the lead anode slime of the above-mentioned factories by wet processing to obtain silver-rich slag, with high gold and silver direct recovery rates (Au 99%, Ag 98%), and can be comprehensively recovered at the same time Antimony, bismuth, copper, lead, the recovery rate antimony, bismuth > 80%, copper) 65% method. The method provided by the present invention includes hydrochloric acid+sulfuric acid mixed acid leaching (or sodium chloride+sulfuric acid medium leaching), hydrolysis of antimony precipitation, hydrolysis of bismuth precipitation, neutralization and replacement of copper precipitation, salt leaching to remove lead, replacement of silver chloride and lead precipitation 1. Obtain silver by smelting silver and electrolysis, several steps of aqua regia dissolving gold and reducing gold, its method step is: 1, lead anode slime is mixed in HCl 2-4mol/L-+H2SO4 2mol/L (or NaCl 3.5-4.5mol/L +H2SO4 2-4mol/L) medium, temperature 80-90°C, ratio of material weight in grams to liquid volume in milliliters is 1:6-10, stirring and leaching for 2-4 hours, antimony, bismuth, and copper are transferred into the solution, Solid-liquid separation; 2. The leaching solution obtained in step 1 is kept at room temperature, stirred and added with tap water until the volume is diluted to 2 times, and a small amount of 10-20% (weight ratio) NaOH solution is added to adjust the pH value of the solution to 1-1.5 to hydrolyze antimony. When the precipitation of SbOCl is completed, the solid-liquid separation is carried out; 3. With the antimony precipitation mother liquor obtained in step 2, at room temperature and under stirring conditions, add 10-20% (weight ratio) NaOH solution to adjust the pH value of the solution to 3.5-4.0 to hydrolyze bismuth, To BiOCl precipitation is complete, solid-liquid separation; 4, with 3 step gained bismuth heavy mother liquors, at temperature, add 10-20% (weight ratio) NaOH solution under stirring condition, adjust solution pH to 6.5-7.0, to Cu ( OH) 2 precipitation is complete; or at room temperature, under stirring conditions, press Cu:Fe=1:1-1.2 (weight ratio), add iron powder to the solution to replace copper, react for 1-2h, and separate solid and liquid; 5 , the leaching slag obtained in step 1, at a pH of 2-4, a temperature of nearly 80°C, and a ratio of the weight of the slag in grams to the volume of the solution in milliliters is 1:13-18, using a NaCl solution with a concentration of 150-200g/L , stirring and leaching for 2 h, so that the lead is completely converted into lead chloride for leaching, and solid-liquid separation; 6. Add lead powder 2 to every 1m3 of the solution at a temperature of 60-70°C under stirring conditions to the lead leaching solution obtained in step 5. -4kg (or iron powder 4-6kg) replaces the silver chloride dissolved in the lead leaching solution, reacts for 1-2h, and separates the solid and liquid; ] Neutralize the replacement solution until the pH is 9-11, precipitate lead and heavy metal ions and sulfate radicals, separate solid and liquid, adjust the pH of the sinking lead mother liquor to 2-4 with hydrochloric acid, and then return to step 5 for initial application; 7. The lead leaching slag gained in step 5 is merged with the silver slag for replacement gained in step 6, and at a temperature of 1050-1100° C., sodium carbonate and sodium nitrate are added for smelting, slag: sodium carbonate: sodium nitrate=1: 1.2-2.5: 0.05-0.1 (weight ratio), time 10-60min, smelting to obtain gold-containing crude silver ingots, and the crude silver is obtained by traditional electrolytic refining method to obtain pure silver; The amount ratio (weight ratio) is 1:3-4. Add in portions, self-heating or post-heating, stirring, so that the gold is completely dissolved, and the silver remains in the slag in the form of silver chloride, and is filtered and separated. The ratio of the amount added is H2C2O4:Au=1-3:1, and the time is 4-6 hours. The sponge gold powder is boiled and washed with 3mol/L HCl, and then washed with water until neutral to obtain pure gold.
本发明的优点是：1、流程以湿法冶金为主，火法湿法结合，流程结构合理；2、金、银直收率高：Au≥99％，Ag≥98％，锑、铋、铜、铅综合利用程度好(回收率Sb、Bi)90％，Cu65％)；3、脱铅置换回收银及母液再生技术，脱铅效率高(＞97％)，银损失小(置换母液含银0.011-0.0044g/l)，能实现脱铅工序溶液闭路作业，有利于环境保护；4、方法适应性强，设备简单，易实现工业化。The advantages of the present invention are: 1. The process is mainly based on hydrometallurgy, combined with fire and wet methods, and the process structure is reasonable; 2. The direct recovery rate of gold and silver is high: Au≥99%, Ag≥98%, antimony, bismuth, The comprehensive utilization of copper and lead is good (recovery rate Sb, Bi) 90%, Cu65%); 3. Deleading substitution recovery silver and mother liquor regeneration technology, high deleading efficiency (> 97%), silver loss is small (replacement mother liquor contains Silver 0.011-0.0044g/l), can realize the closed-circuit operation of the deleading process solution, which is beneficial to environmental protection; 4, the method has strong adaptability, simple equipment, and easy industrialization.
实施例1、铅阳极泥1578.35Kg，主要成份(％)，Au 0.0085-0.0102，Ag 22.416，Pb 14.27，Bi 0.046，Cu 9.78，Sb 30.11。采用HCl 3mol/L+H2SO4 1mol/L介质，温度80-85℃，铅泥重量(Kg)∶液体体积(L)＝1∶10，搅拌浸出2h，固液分离，浸出液按2步骤所列水解锑条件沉锑；由于铅阳极泥含铋很低不予回收，沉锑母液直接用废铁屑置换沉铜得粗铜粉；酸浸渣于80℃，用200g/L NaCI溶液，按渣重量(Kg)∶液体体积(L)＝1∶15，搅拌浸铅2h，固液分离，浸铅液在60℃温度下，用6步骤所列铅置换氯化银条件回收少量银，置换母液经石灰沉铅，工业盐酸回调PH后返回浸铅复用，浸铅渣按配料比：渣∶Na2CO3∶NaNO3＝1∶1.5∶0.1(重量比)、温度1100℃，进行熔炼得到金银合金锭389.08kg，品位：Au 408-471g/t，Ag 91.28-95.39％；回收率：Au＞99％，Ag 99％；产出的副产品品位，锑渣Sb 39.36-59.76％，粗铜粉Cu 31.59％，铅渣Pb 31.17-42.72％。Embodiment 1, lead anode slime 1578.35Kg, main component (%), Au 0.0085-0.0102, Ag 22.416, Pb 14.27, Bi 0.046, Cu 9.78, Sb 30.11. Use HCl 3mol/L+H2SO4 1mol/L medium, temperature 80-85°C, lead slime weight (Kg): liquid volume (L) = 1:10, stir for leaching for 2 hours, separate solid and liquid, and hydrolyze the leachate as listed in 2 steps Precipitation of antimony under antimony conditions; due to the low content of bismuth in the lead anode slime, it is not recycled, and the mother liquor of antimony deposition is directly replaced with scrap iron scraps to obtain coarse copper powder; the acid leaching residue is at 80 °C, using 200g/L NaCI solution, according to the weight of the slag (Kg): liquid volume (L)=1: 15, stirring and soaking lead 2h, solid-liquid separation, soaking lead liquid is at 60 ℃ of temperature, and reclaims a small amount of silver with the lead replacement silver chloride condition listed in 6 steps, replaces mother liquor through Lime sinks lead, industrial hydrochloric acid adjusts the PH and returns to lead leaching for reuse. The lead leaching slag is smelted according to the ingredient ratio: slag: Na2CO3: NaNO3 = 1: 1.5: 0.1 (weight ratio), and the temperature is 1100 ° C. Gold and silver alloy ingots 389.08 kg, grade: Au 408-471g/t, Ag 91.28-95.39%; recovery rate: Au>99%, Ag 99%; grade of by-products produced, antimony slag Sb 39.36-59.76%, blister copper powder Cu 31.59%, Lead slag Pb 31.17-42.72%.
实施例2铅阳极泥1000g，主要成分为(％)：Au 0.133，Ag 10.946，Pb 12.40，Bi 5.78，Cu 2.29，Sb 42.38。采用NaCl 4.3mol/L+H2SO4 4mol/L介质，温度80℃，铅泥重量(g)∶液体体积(ml)＝1∶8，搅拌浸出2h，固液分离，浸出液按2步骤、3步骤、4步骤所列条件水解沉锑，水解沉铋及铁粉置换沉铜，浸出渣用NaCl 200g/L溶液，温度80-85℃，按渣重量(g)∶液体体积(ml)＝1∶15，搅拌浸铅2h，固液分离；浸铅液在搅拌条件下，按每1L加入2.2g铅粉，65℃温度下，置换反应2h，固液分离，浸铅渣236.5g，加入碳酸钠425g，硝酸钠12g，于1100℃温度熔炼10min，得到粗银锭110g，品位Au 1.214％，Ag 97.29％，加上铅置换银渣18g中含银量0.58g，回收率：Au 99.86％，Ag 98.35％；产出副产品锑渣805g，含Sb 50.11％，回收率95.28％，铋渣199g，含Bi 59.50％，回收率99％；粗铜粉19.5g，含Cu 80.07％，回收率68.16％，铅渣122.9g，Pb品位29.89％。Example 2 Lead anode slime 1000g, the main components are (%): Au 0.133, Ag 10.946, Pb 12.40, Bi 5.78, Cu 2.29, Sb 42.38. Use NaCl 4.3mol/L+H2SO4 4mol/L medium, temperature 80°C, lead slime weight (g): liquid volume (ml) = 1:8, stirring and leaching for 2 hours, solid-liquid separation, leaching solution according to 2 steps, 3 steps, The conditions listed in step 4 are to hydrolyze antimony, hydrolyze bismuth and replace iron powder to replace copper, use NaCl 200g/L solution for leaching slag, temperature 80-85°C, weight of slag (g): liquid volume (ml) = 1:15 , stirring and soaking lead for 2 hours, solid-liquid separation; under the condition of stirring, add 2.2g of lead powder per 1L of the lead soaking solution, at 65°C, replace the reaction for 2 hours, separate the solid and liquid, add 236.5g of lead slag, add 425g of sodium carbonate , 12g of sodium nitrate, smelted at 1100°C for 10min to obtain 110g of crude silver ingots, with grades of Au 1.214%, Ag 97.29%, and silver content of 0.58g in 18g of lead-substituted silver slag, recovery rate: Au 99.86%, Ag 98.35% Output by-product antimony slag 805g, containing Sb 50.11%, recovery 95.28%, bismuth slag 199g, containing Bi 59.50%, recovery 99%; Blister copper powder 19.5g, containing Cu 80.07%, recovery 68.16%, lead slag 122.9g, Pb grade 29.89%.
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