WO2014154160A1 - 一种硫酸铅湿法炼铅工艺 - Google Patents
一种硫酸铅湿法炼铅工艺 Download PDFInfo
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- WO2014154160A1 WO2014154160A1 PCT/CN2014/074192 CN2014074192W WO2014154160A1 WO 2014154160 A1 WO2014154160 A1 WO 2014154160A1 CN 2014074192 W CN2014074192 W CN 2014074192W WO 2014154160 A1 WO2014154160 A1 WO 2014154160A1
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- zinc
- lead
- extraction
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- stripping
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B13/00—Obtaining lead
- C22B13/04—Obtaining lead by wet processes
- C22B13/045—Recovery from waste materials
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/30—Obtaining zinc or zinc oxide from metallic residues or scraps
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
- C22B3/10—Hydrochloric acid, other halogenated acids or salts thereof
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
- C22B3/44—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes
- C22B3/46—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes by substitution, e.g. by cementation
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- 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
Definitions
- the invention belongs to a hydrometallurgical process technology, and relates to a technology for producing lead by using a full wet metallurgical process for lead slag or lead storage lead produced by wet zinc smelting.
- the wet zinc smelting raw material contains 1-4% lead.
- the lead is enriched in the high leaching slag during the zinc leaching process.
- the form of lead is almost all PbSO 4 .
- the leaching slag is treated by the fire method, lead and zinc are simultaneously volatilized into the soot, and after further treatment of the leaching of the smog, the lead is concentrated in the leaching slag, and the form thereof is PbSO 4 .
- the above-mentioned various PbSO 4 slag materials generally contain 10-40% lead.
- fire smelting is usually used, which not only consumes high energy, but also generates flue gas accompanied by lead dust and lead.
- Gas phase pollution such as smoke, SO 2 fumes, and dioxins.
- the lead storage battery When the used lead storage battery is recycled, the lead storage battery is disassembled and crushed, and the lead sludge is sorted out.
- the main components are PbSO 4 and PbO.
- lead-nitride recovery in the world is mainly based on pyrometallurgical smelting, and there is also high energy consumption. Air pollution problems.
- the object of the present invention is to provide a lead sulfate wet process lead-smelting process, which uses lead sulfate slag (mud) as a raw material to directly produce metal lead, has the characteristics of low energy consumption, low cost, easy to popularize and apply, and completely eliminates lead smoke. Flue gas pollution such as gas, lead dust and SO 2 .
- the solution adopted for achieving the object of the present invention is: using CaCl 2 + NaCl as a leaching solvent, disassembling the high-impregnated slag of zinc-containing wet zinc smelting, the lead slag after leaching zinc of lead-containing zinc oxide or the lead-acid battery
- the lead mud is leached, the lead solution is replaced with zinc, the zinc is recovered after the replacement, the recovered zinc is returned to the lead replacement process, and the CaCl 2 + NaCl solution after the recovery of zinc is recycled to the lead sulfate chloride salt leaching process, wherein 1Ca 2 + molar ratio ⁇ SO 4 2- molar ratio in the material; 2
- the ratio of leaching solution of NaCl and CaCl 2 should be adjusted according to the real-time monitoring data in the operation to ensure that the solution contains Ca 2+ of 3 ⁇ 4g/L and Pb 2+ after leaching.
- the temperature of the leachate is controlled at 18 °C ⁇ 90 °C, the pH of the leachate is 4 ⁇ 5; the liquid after soaking should be cleaned with sponge lead to remove trace impurities of Ag, Cu and Bi in the solution; the zinc used for the replacement is metal zinc sheet and zinc powder. Zinc particles and zinc balls are the best zinc sheets precipitated by electrolytic zinc.
- the replacement temperature is 18 ° C to 90 ° C, and the amount of zinc replacement is 105 to 110% of the theoretical amount.
- the leaching solvent used in the present invention is CaCl 2 + NaCl, and the leaching process is carried out by pressing the reaction:
- the ratio of NaCl to CaCl 2 can be adjusted as required.
- CaCl 2 has two functions: one is to thoroughly carry out the above reaction, and the Ca 2+ molar ratio ⁇ the SO 4 2- molar ratio in the material, so that all of the SO 4 2- in the material during the leaching process is more stable than the lead sulfate.
- the CaSO 4 • 2H 2 O precipitates to minimize SO 4 2- in the solution to facilitate subsequent processing of the leachate.
- the second is to increase the solubility of PbCl 2 in the solution during the reaction.
- the role of NaCl is mainly to provide sufficient Cl - to dissolve the PbCl 2 produced by the reaction.
- the solubility of PbCl 2 in the solution is related to the solution Cl - concentration and CaCl 2 concentration, and the solution temperature.
- PbSO 4 so that the reaction is thoroughly, quickly, after leach solution containing Ca 2+ ⁇ 3 ⁇ 4g / L; leaching solution to achieve the desired concentration of Pb 2+ 20g / L, solution Cl - concentration ⁇ 150g / L;
- the temperature is selected according to the Cl - and CaCl 2 contents, and when Cl - and CaCl 2 are high, the temperature is suitably low; when Cl - and CaCl 2 are low, the temperature must be raised.
- the solution ratio is different, and the temperature is between 18 ° C and 90 ° C.
- the leaching solution solid ratio is determined according to the amount of Pb contained in the material, and the leaching solution is controlled to contain Pb 2+ 10 ⁇ 30g/L, and the leaching conditions have greater flexibility.
- the leaching was carried out in a stirred vessel, the leaching rate of Pb was 95%, and the pH of the leaching solution was 4-5.
- the lead slag itself is acidic and does not require additional acid.
- a small amount of hydrochloric acid should be added, and the leaching end point pH 4-5 should be adjusted with lime milk.
- the leaching solution is purified by the solution, and the purification is carried out in a stirred tank. 1 g/L of sponge lead is added to the tank to remove trace impurities such as Ag, Cu, Bi, etc., and the solution is replaced by zinc after purification. Lead, the replacement reaction process is:
- Lead is replaced by metal zinc, which can be made into different physical forms, namely zinc powder, zinc flakes, zinc flakes, and zinc balls.
- the zinc flakes deposited by electrolytic zinc are the most reasonable and require no additional processing.
- the displacement is carried out in a displacement reactor in which a chlorine salt leaching solution containing PbCl 2 and metallic zinc are added, the displacement reaction proceeds rapidly, and the solution is allowed to stand in the displacement reactor for ⁇ 5 minutes, and is replaced by PbCl 2 in the displacement solution.
- Metal lead When replacing lead, it is not necessary to specifically heat up and cool down, and it can be carried out at the natural temperature of the leachate, that is, 18 ° C to 90 ° C.
- the amount of zinc used in the replacement is 0.3165 t/t lead according to the theoretical amount, and the actual amount is 105-110% of the theoretical amount.
- the lead contained in the solution contained lead Pb ⁇ 50 mg / L.
- the first solution is to directly extract zinc from the displaced liquid, back stripping, electrolysis, and electrolytic zinc sheet for replacing lead.
- the raffinate is returned to leach the lead sulfate and recycled.
- the second scheme is to neutralize the precipitated zinc, and the zinc solution is returned to the lead sulfate leaching and recycled.
- the precipitated Zn(OH) 2 is dissolved by sulfuric acid (raffinate) ⁇ zinc is extracted ⁇ zinc is stripped ⁇ electrolytic zinc is used, and zinc flakes are used for replacing lead.
- the first scheme of zinc recovery is as follows: the zinc is directly extracted after the replacement, and the extractant is P204, which is diisooctyl phosphate.
- the process of zinc recovery is saponification of extractant ⁇ extraction of zinc ⁇ washing ⁇ back extraction ⁇ electrolytic zinc. The process is as follows:
- the extractant contains about 5 to 12 g/L of Ca for extracting zinc.
- the extraction temperature is 20-40 ° C, and Ca is transferred to the raffinate to form CaCl 2 during the extraction process.
- the raffinate is returned to the leaching of PbSO 4 slag and recycled.
- the zinc electrolysis waste liquid contains H 2 SO 4 150-180g/L, contains Zn 40 ⁇ 50g/L, and is mixed with the supported zinc organic phase to strip the zinc into the stripping solution.
- the stripping reaction is based on Zn 90 after stripping. Controlled by ⁇ 120g/L, the stripping can be carried out in a countercurrent flow of one or two stages, and the organic phase contains Zn ⁇ 0.5g/L.
- the reaction of stripping zinc is:
- the stripping mixing time is 0.5 to 10 minutes, the stripping temperature is 20 to 40 ° C, and the organic phase after stripping is returned to saponification and recycled.
- Stripping solution degreasing The stripping agent is usually entrained or dissolved in the stripping solution, usually about 10PPm. When P204>1PPm, it will affect the zinc electrolysis, causing the current efficiency to drop. With air flotation-activated carbon adsorption, P204 can be reduced to ⁇ 1PPm for electrolysis, and the electrolysis process is the same as conventional zinc electrolysis.
- the zinc electrolysis reaction is:
- the regenerated H 2 SO 4 is returned to the stripped zinc, recycled, and the zinc flakes are used to replace the lead.
- excess leached zinc may be recovered, and the excess zinc ingot may be sold as a product.
- the second scheme for zinc recovery includes neutralization of precipitated zinc, Zn(OH) 2 dissolution, zinc extraction, stripping and electrolysis.
- the zinc-plated solution is returned to the leaching of PbSO 4 and recycled.
- the Zn(OH) 2 precipitate was separated, and the entrained mother liquor was washed away with water, and the precipitate contained 55 to 64% of Zn (dry basis).
- Zn (OH) 2 were dissolved: dissolving Zn (OH) with dilute sulfuric acid (raffinate) 2, the control leachate containing Zn 20 ⁇ 30g / L, in a stirred tank, at room temperature. After dissolving the zinc, the solution was filtered and subjected to zinc extraction.
- the zinc dissolution reaction is:
- the raffinate is returned to Zn(OH) 2 and dissolved for recycling.
- the extraction series is carried out in a countercurrent flow of 3 to 4 stages.
- the extracted organic phase contains 10-15 g/L of zinc, and is stripped with zinc electrolysis waste liquid.
- the stripping process is the same as that of Scheme 1. After stripping, the organic phase is returned to the zinc for extraction and recycled.
- the invention relates to a hydrometallurgical process for producing lead by using PbSO 4 mud (slag) as raw material.
- PbSO 4 is leached by a mature chloride salt leaching process to form a PbCl 2 chloride salt solution, while the sulfate is converted to harmless gypsum.
- PbCl 2 is reduced to metal lead by means of metal zinc, and then extracted from the displaced liquid by zinc solvent ⁇ electrolytic zinc to achieve zinc circulation and return to replace lead.
- the calcium required for the leaching process is supplemented during the recovery of zinc. Has the following characteristics:
- the first scheme of zinc recovery according to the present invention is (1) + (2) + (2) + (4) + (5) + (6), and the total process reaction is:
- the second reaction of zinc recovery is (1) + (2) + (3) + (7) + (8) + (9) + (5) + (6), and the total reaction is still (10).
- the process consumes Ca(OH) 2 and direct current, and the other processes are intermediate conversion processes.
- Leaching of PbSO 4 relies on the reaction of CaCl 2 and PbSO 4 to convert PbSO 4 into PbCl 2 and CaSO 4 • 2H 2 O.
- the reaction is thorough and rapid, and the Cl - total of the leachate is controlled according to the amount of PbCl 2 dissolved. the amount.
- the reaction produces CaSO 4 • 2H 2 O, which is environmentally friendly, ie gypsum.
- All process solutions of the present invention are closed loops. After the leachate is replaced by Pb and then extracted and zinc is extracted, CaCl 2 is recirculated to leaching PbSO 4 . In the second scheme, after replacing Pb, the zinc is neutralized, CaCl 2 is regenerated and returned to PbSO 4 leaching, and the zinc raffinate and the solution Zn(OH) 2 form a closed loop during the precipitation-extraction-back extraction electrowinning process of zinc; In the extraction-back extraction zinc-electrowinning process, the electrolytic waste liquid forms a closed loop with the stripping zinc; the organic phase in the extraction process forms an extraction-back extraction closed loop. No wastewater discharge.
- the invention has obvious environmental friendliness advantages, the whole process is a full wet process, the sulfate in the material is converted into safe and harmless gypsum (CaSO 4 • 2H 2 O), no SO 2 flue gas emission, lead-free Dust and lead smoke emissions.
- the present invention produces lead from PbSO 4 sludge, which is low in cost, and the material consumed in the process is Ca(OH) 2 and direct current consumption.
- Ca(OH) 2 is cheap and readily available lime, and direct current is zinc electrolysis power consumption.
- the replacement of 1 ton of lead theoretically requires 0.3165 tons of zinc zinc, and the actual consumption is 0.34 to 0.35 tons.
- the direct current consumption of zinc electrolysis The DC consumption of lead and lead is 1020 ⁇ 1050kwh/ton.
- the comprehensive cost is 3000 ⁇ 3200 yuan/ton according to the current price level.
- the cost of the lead unit is 0.150 ⁇ 1120 yuan/ton lead.
- the production cost is very low. For other processes.
- the invention has the advantages of environmental friendliness, low cost, easy control, and easy realization of large-scale production.
- Figure 1 is a schematic diagram of the process flow of the present invention.
- FIG. 2 is a schematic diagram of a process flow of a first solution combining zinc recovery according to the present invention.
- FIG. 3 is a schematic diagram of a process flow of a second solution combining zinc recovery according to the present invention.
- the slag contains 21.89% lead, all of which is lead sulfate, containing Zn 3.5% and containing S10.2%.
- the leachate was a chloride salt solution, NaCl 280 g/L, and Ca 2+ 13 g/L.
- Leaching conditions 500 g (dry weight) of slag was added to the leachate 5 L, and leached under stirring.
- the liquid-solid ratio was 10:1, the temperature was 80 ° C, the leaching time was 1.5 hours, and the leaching end point was pH 4.5.
- the leaching residue contained Pb 0.9%, the leaching slag rate was 96%, and the leaching solution contained Pb 21.1 g/L.
- the leaching solution + washing water is 5.32L, containing Pb 16.0g/L, and the Pb leaching rate is 93.15%.
- Lead-acid battery lead leaching a lead smelter battery recycling workshop, the disassembled lead mud, containing lead 76.04% (dry basis), taking lead paste 100g, leachate 2L, leaching solution NaCl 320g / L, HCl 20g /L, leaching temperature 85 ° C, leaching time 2 hours, leaching end point pH 2.0, adding Ca(OH) 2 to adjust the pH to 4.5. Filtration filtrate + washing water 2.1L, containing Pb 34.8g / L, lead leaching rate 96 %.
- the leaching results are as follows:
- PbCl 2 chlorinated leachate the composition is as follows:
- Chlorine salt solution composition of PbCl 2 Chlorine salt solution composition of PbCl 2 :
- the pellet was covered with 320 g of NaOH and melted at 500 ° C to obtain a lead ingot.
- the extractant used industrial P204, with 260 # kerosene as the diluent, and the volume ratio was 20%, and the theoretical saturation capacity was Zn 2+ 17.8g/L and Ca 2+ 10.9g/L, respectively.
- Organic phase saponification CaCl 2 is added to the saponification solution, the solution contains Ca 2+ 9.83 g / L, and the phase separation is carried out with a 20% P204 solution in a 1:1 ratio with the saponification solution for 4 minutes.
- the saponified solution contains Ca 2+ 0.9. g/L, pH 4.5; organic phase containing Ca 2+ 8.9g / L, for zinc extraction, saponification, adding Ca (OH) 2 to add Ca 2+ to about 10g / L, recycling.
- the organic phase and the aqueous phase were analyzed.
- the organic phase contained Zn 14.66 g/L and Ca 0.052 g/L.
- the raffinate contains Ca 10.5 g/L and Zn 4.2 g/L, and returns to the leaching of PbSO 4 slag.
- Stripping zinc After extracting zinc, the organic phase is back-extracted with a zinc-containing sulfuric acid solution.
- the zinc and acid content of the stripping solution is the same as that of the zinc electrolytic acid solution.
- reagent stripping solution distilled water.
- the liquid composition meets the requirements of zinc electrolysis.
- the liquid can be subjected to zinc electrolysis after being subjected to air flotation-adsorption by activated carbon.
- Zinc electrolysis produces zinc flakes according to the zinc extraction-electrolytic zinc process of the enterprise.
- Neutralizing zinc is the second option for zinc recovery.
- Liquid component g/L after lead replacement Zn Pb Ca 14.1 0.02 3.2
- the replacement liquid Take 5L of the replacement liquid, add powdered Ca(OH) 2 , carry out in a stirred tank, neutralize to pH 7.0, produce 119.1g of Zn(OH) 2 , 59.2% of Zn, and 5mg of Zn after immersion zinc /L. Containing Ca 2+ 12.3g / L, the zinc immersion liquid can return to PbSO 4 leaching.
- Dissolution of Zn(OH) 2 In order to meet the needs of subsequent zinc recovery, the zinc raffinate produced by our company's zinc extraction is used for dissolution.
- Raffinate component g/L Zn H 2 SO 4 7.5 twenty one
- the dissolved liquid meets the zinc extraction requirements, and the zinc extract is produced by a zinc extraction-electrolytic zinc process.
Abstract
一种硫酸铅湿法炼铅工艺,该工艺以CaCl2 + NaCl为浸出溶剂,对含有硫酸铅的湿法炼锌的高浸渣、含铅氧化锌浸出锌后的铅渣或铅酸蓄电池拆解后的铅泥浸出,浸后液用锌置换铅,置换后液回收锌,回收的锌返回置换铅工序,回收锌后的含Cl-液循环到硫酸铅氯盐浸出工序。该工艺具有低能耗、低成本、易于推广应用的特点,彻底消除了铅烟气、铅尘和SO2等烟气污染。
Description
本发明属于湿法冶金工艺技术,涉及对湿法炼锌附产的硫酸铅渣或铅蓄电池铅泥采用全湿法冶金工艺生产铅的技术。
湿法炼锌原料中含1~4%铅,锌焙砂采用高酸浸出针铁矿法沉铁或铁矾法沉铁工艺时,在锌浸出过程中,铅富集在高浸渣中,其中的铅的形态几乎都是PbSO4。采用火法处理浸出渣时,铅和锌同时挥发进入烟尘,进一步处理烟尘浸出锌后,铅集中于浸出渣,其形态均为PbSO4。
其它含锌物料如钢铁厂烟尘、炼铜厂烟尘等经火法还原挥发处理后的氧化锌同样含有铅,进一步处理浸出锌后,铅也集中在浸出渣中,其形态均为PbSO4。
上述各种含PbSO4渣物料含铅量一般为10~40%,为了从上述物料中回收铅,通常采用火法熔炼,不但能耗高,冶炼过程产生的烟气同时伴随着铅尘、铅烟气、SO2烟气、二恶英等气相污染。
废旧铅蓄电池回收时,铅蓄电池经过拆解、破碎,分选出的铅泥,主要成分为PbSO4和PbO,目前世界各国处理铅泥回收铅以火法冶炼为主,同样存在能耗高和空气环境污染问题。
国内外学者也研究了铅蓄电池铅泥的湿法处理工艺,普遍采用转化法将PbSO4、PbO转化为PbCO3或PbCl2。碳酸盐转化法将PbSO4加碳酸钠或碳酸铵转化脱硫,同时PbSO4转化为PbCO3,用硼氟酸或硅氟酸溶解碳酸铅以硼氟酸铅或硅氟酸铅电解产出金属铅。氯化法用盐酸和氯盐浸出铅生成氯化铅,采用隔膜电解产出金属铅。这些电解方法都存在铅电解复杂、困难等问题。因此,至今还没有规模化生产实践。
在世界上环保要求越来越高,环境压力日益增大的形势下,PbSO4物料的湿法处理是当前世界冶金界和环境保护界研究的热点之一,但都存在各种不同问题,没有实现规模化生产。
本发明的目的是提供一种硫酸铅湿法炼铅工艺,以硫酸铅渣(泥)为原料,直接产出金属铅,具有低能耗、低成本、易于推广应用的特点,彻底消除了铅烟气、铅尘和SO2等烟气污染。
实现本发明目的所采取的方案是:以CaCl2 +
NaCl为浸出溶剂,对含有硫酸铅的湿法炼锌的高浸渣、含铅氧化锌浸出锌后的铅渣或铅酸蓄电池拆解后的铅泥浸出,浸后液用锌置换铅,置换后液回收锌,回收的锌返回置换铅工序,回收锌后的含CaCl2
+
NaCl液循环到硫酸铅氯盐浸出工序,其中,①Ca2+摩尔比≥物料中SO4
2-摩尔比;②NaCl和CaCl2的浸出液配比应根据操作中实时监测数据调整,保证浸出后溶液含Ca2+为3~4g/L、含Pb2+
为10~30g/L,浸出液中Cl-浓度≥150g/L;③锌回收有两种方案,即a.置换后液采用P204为萃取剂,经过萃取剂皂化→萃取锌→洗杂→反萃→电解锌一系列工序,b.
用Ca (OH)2中和沉锌,再经过Zn(OH)2溶解,锌萃取,反萃和电解锌四个工序。
浸出液温度控制在18℃~90℃,浸出液的pH=4~5;浸后液应采用海绵铅净化,除去溶液中的Ag、Cu、Bi微量杂质;置换所用的锌为金属锌片、锌粉、锌粒、锌球,以电解锌析出的锌片最佳,置换温度为18℃~90℃,置换锌用量为理论用量的105~110%。
以下对发明的几个重要工序进行详述。
一、浸出
本发明采用的浸出溶剂为CaCl2 + NaCl,浸出过程按下反应进行:
PbSO4 + CaCl2 +
2H2O = CaSO4•2H2O↓+ PbCl2 ……(1)
NaCl和CaCl2的配比可按照要求调整。CaCl2有两种功能:一是使上述反应进行彻底,Ca2+摩尔比≥物料中SO4
2-摩尔比,使在浸出过程中物料中的SO4
2-全部转为比硫酸铅更稳定的CaSO4•2H2O沉淀,使溶液中SO4
2-到最少,以利于浸出液的后续处理。二是增大反应过程的PbCl2在溶液中的溶解度。NaCl的作用主要是提供足够的Cl-,溶解反应生成的PbCl2。PbCl2在溶液中的溶解度与溶液Cl-浓度和CaCl2浓度、溶液温度有关。随着溶液Cl-浓度升高,CaCl2含量升高,温度升高,PbCl2的溶解度升高。为使PbSO4反应能彻底、快速进行,浸出后溶液含Ca2+≥3~4g/L;为达到浸出液中Pb2+浓度达到所需的20g/L,溶液Cl-浓度≥150g/L;温度根据Cl-和CaCl2含量选取,Cl-和CaCl2高时,可适当低温;Cl-和CaCl2低时,则必须提高温度。溶液配比不同,温度在18℃~90℃之间。浸出液固比按物料含Pb量确定,控制在浸出液含Pb2+
10~30g/L,浸出条件有较大灵活性。
浸出在带搅拌容器中进行,Pb的浸出率95%,浸出液pH4~5。在处理锌冶炼的铅渣时,铅渣本身为酸性,不需另外加酸。处理蓄电池铅泥时需加入少量盐酸,用石灰乳调整浸出终点pH4~5。
二、溶液净化除杂及铅置换
将浸出矿浆进行液固分离后,浸出液进行溶液净化,净化在搅拌槽内进行,在槽内加入1g/L海绵铅,除去溶液中的Ag、Cu、Bi等微量杂质,净化后溶液用锌置换铅,置换反应过程为:
PbCl2 + Zn = Pb + ZnCl2
……(2)
用金属锌置换铅,金属锌可以制成不同物理形态,即锌粉、锌片、锌粒、锌球均可。以电解锌析出的锌片最为合理,不需额外加工。
置换在置换反应器中进行,反应器中加入含PbCl2的氯盐浸出液和金属锌,置换反应进行很快,溶液在置换反应器中停留时间≥5分钟,通过置换溶液中的PbCl2被置换成金属铅。置换铅时不需专门升温、降温,以浸出液自然温度进行,即18℃~90℃均可。置换时锌用量按理论量为0.3165t/t铅,实际用量为理论用量105~110%。
置换过程,置换后液含铅Pb<50mg/L。
将置换金属铅取出,经漂洗后压团,
以团块量10~20%的烧碱(NaOH)覆盖,在500℃温度下熔化得到金属铅,铅含量>99%。铅品位杂质含量由溶液净化程度确定。
三、锌回收
锌的回收有两种方案。
第一方案是从置换后液中直接萃取锌→反萃→电解,电解锌片用于置换铅。萃余液返回浸出硫酸铅,循环使用。
第二方案是中和沉淀锌,沉锌后液返回硫酸铅浸出,循环使用。沉淀的Zn(OH)2经硫酸(萃余液)溶解→萃取锌→反萃锌→电解锌,
锌片用于置换铅。
锌回收第一方案
锌回收第一方案具体如下:置换后液直接萃取锌,萃取剂采用P204即磷酸二异辛醋。锌回收的过程为萃取剂皂化→萃取锌→洗杂→反萃→电解锌。过程如下:
(1)皂化:用Ca(OH)2对P204进行皂化,皂化液含Ca(OH)2,P204以260#溶剂油按体积比P204:煤油=20~30%,用配制好的P204和皂化剂混合均匀1~2分钟,静置沉清分相,皂化反应按下化学反应进行:
2HR(有) + Ca(OH)2(水)
=CaR2(有) + 2H2O) ……(3)
皂化后萃取剂中含Ca约5~12g/L,用于萃取锌。
(2)萃取锌:皂化后的P204与置换后液混合,澄清分相,混合时间1~5分钟。萃取反应为:
CaR2(有相) +ZnCl2(水相)
=ZnR2(有相) + CaCl2(水相) ……(4)
采用3~5级逆流萃取,萃取温度20~40℃,萃取过程中Ca转入萃余液形成CaCl2。
萃余液返回PbSO4渣的浸出,循环使用。
由于水相与有机相分离时,有机相中不可避免夹带少量萃余液,用水洗涤除去。
(3)反萃锌:萃锌后有机相用锌电解废液进行反萃。锌电解废液含H2SO4 150
~180g/L,含Zn 40~50g/L,与负载锌有机相混合,将锌反萃进入反萃液,反萃相比按照反萃后含Zn
90~120g/L进行控制,反萃可按一级或二级逆流进行,以有机相含Zn<0.5g/L为准。反萃锌的反应为:
ZnR2(有) +H2SO4(水)
= ZnSO4(水) + 2HR(有) ……(5)
反萃混合时间0.5~10分钟,反萃温度20~40℃,反萃后的有机相返回皂化,循环使用。
反萃液除油:反萃液中通常夹带或溶解微量萃取剂,通常为10PPm左右,当P204>1PPm时对锌电解会产生影响,引起电流效率下降。采用气浮—活性炭吸附,P204可降至<1PPm,进行电解,电解过程与常规锌电解相同。
锌电解反应为:
ZnSO4 + H2O+ 2e = Zn +
H2SO4+ 1/2O2 ……(6)
再生的H2SO4返回反萃锌,循环使用,析出锌片用于置换铅。当处理锌浸出后的PbSO4渣时,可附带回收多余浸出的锌,多余的锌铸锭后作为产品销售。
锌回收第二方案
锌回收第二方案包括中和沉淀锌,Zn(OH)2溶解,锌萃取,反萃和电解五个过程。
(1)中和沉锌:置换后液含NaCl +CaCl2 +
ZnCl2,用Ca
(OH)2中和沉锌,在搅拌槽内加入Ca(OH)2,pH升高至7~9,使锌沉淀,中和沉锌反应为:
ZnCl2+ Ca(OH)2 =
Zn(OH)2↓+ CaCl2 ……(7)
进行液固分离后,沉锌后液返回PbSO4的浸出,循环使用。分离出Zn(OH)2沉淀物,再用水洗除夹带的母液,沉淀物含Zn
55~64%(干量计)。
(2)Zn(OH)2溶解:用稀硫酸(萃余液)溶解Zn(OH)2,控制浸出液含Zn
20~30g/L,在搅拌槽内进行,常温。溶解锌后溶液过滤后进行锌萃取。锌溶解反应为:
Zn(OH)2↓+ H2SO4 =
ZnSO4 + 2H2O ……(8)
(3)萃取锌:采用P204,以260#溶剂煤油为稀释剂,配制成P204含量为20~40%的有机相,用有机相与浸出液搅拌混合,使锌转入有机相。萃取反应为:
3HR(有) + ZnSO4(水)=
ZnR2•HR(有) + H2SO4(水) ……(9)
萃余液返回Zn(OH)2溶解,循环使用。萃取级数3~4级逆流进行。有机相与水相比O/A =
1/2~2/1,萃取混合时间0.5~5分钟,萃取温度20~30℃。
(4)反萃锌:经萃取后的有机相含锌10~15g/L,用锌电解废液进行反萃,反萃过程与方案1相同。反萃后有机相返回萃取锌,循环使用。
(5)反萃液除油电解锌
:过程与方案1相同,电解废液返回反萃锌,循环使用,析出锌片用于氯化铅置换。
本发明为以PbSO4泥(渣)为原料生产铅的湿法冶金工艺。用成熟的氯盐浸出工艺浸出PbSO4,生成PbCl2氯盐溶液,同时硫酸盐转化为无害的石膏。借助金属锌将PbCl2还原为金属铅,再从置换后液中用锌溶剂萃取→电解锌,实现锌的循环,返回置换铅。浸出过程所需的钙在回收锌的过程中补充。具有以下特点:
(1)按本发明的锌回收第一方案反应为(1)+(2)+(2)+(4)+(5)+(6),工艺总反应为:
PbSO4 + Ca(OH)2 +
H2O + 2e = Pb + CaSO4•2H2O↓+ 1/2O2
……(10)
锌回收第二方案反应为(1)+(2)+(3)+(7)+(8)+(9)+(5)+(6),总反应仍然为(10)。
过程中消耗的是Ca(OH)2和直流电,其它过程均为中间转化过程。
(2)PbSO4的浸出依靠CaCl2和PbSO4的反应将PbSO4转化为PbCl2和CaSO4•2H2O,反应进行彻底、快速,按照所需溶解的PbCl2量控制浸出液的Cl-总量。反应生成对环境无害的CaSO4•2H2O即石膏。
(3)生成CaSO4•2H2O所需的原料为廉价易得的石灰,石灰的补充在锌回收过程中完成。方案1是在锌萃取时皂化和锌萃取过程中用Ca(OH)2补充所需的钙,通过锌的萃取过程完成。方案2是置换铅后液中和沉淀锌时以Ca(OH)2形态补充。
(4)用锌置换铅,PbCl2转为金属铅借助锌的置换实现。置换用的锌通过锌回收工艺反复循环使用,采用低成本、高回收率的锌萃取→反萃→电解法,实现循环利用,同时可附带回收原料中的部分溶解的锌。按理论计量置换一顿铅需3.16吨锌,实际消耗为理论量105~110%,用成熟、可靠的锌回收电解工艺取代了复杂昂贵的铅电积,使湿法炼铅可以实现规模产业化。
(5)本发明所有工艺溶液均为闭路循环。浸出液置换Pb后再萃取提锌后CaCl2得到再生循环至浸出PbSO4。第二方案在置换Pb后,中和沉锌,CaCl2得到再生返回至PbSO4浸出,锌的沉淀—萃取—反萃电积过程中锌萃余液与溶液Zn(OH)2形成闭路循环;萃取—反萃锌—电积锌过程中电解废液与反萃锌形成闭路循环;萃取过程有机相形成萃取—反萃闭路循环。无废水排放。
(6)本发明具有明显的环境友好优势,整个过程为全湿法流程,物料中的硫酸盐转化成安全无害的石膏(CaSO4•2H2O),无SO2烟气排放,无铅尘、铅烟气排放。
(7)本发明由PbSO4渣泥生产铅,成本低廉,过程中消耗的物料为Ca(OH)2和直流电消耗。Ca(OH)2为价廉易得的石灰,直流电为锌电解电耗,按本发明置换1吨铅理论上需锌片0.3165吨锌,实际消耗为0.34~0.35吨,按锌电解直流电耗,折合吨铅直流电耗1020~1050kwh/吨,锌回收过程中按当前物价水平综合成本3000~3200元/吨,折合分摊到吨铅单元成本为铅1050~1120元/吨铅,生产成本大幅度低于其它工艺。
本发明具有环境友好,成本低廉,易于控制,易于实现规模化生产等优点。
图1为本发明的工艺流程简图。
图2为本发明结合锌回收第一方案工艺流程简图。
图3为本发明结合锌回收第二方案工艺流程简图。
实例1:浸出
某企业挥发窑氧化锌,浸出回收锌、铟后的硫酸铅渣。渣含铅21.89%,全部为硫酸铅,含Zn3.5%、含S10.2%。
浸出液为氯盐溶液,NaCl 280g/L,Ca2+
13g/L。浸出条件:500g(干重)渣加入浸出液5L,在搅拌条件下浸出,液固比10:1,温度80℃,浸出时间1.5小时,浸出终点pH4.5。
浸出渣含Pb 0.9%,浸出渣率96%,浸出液含Pb 21.1 g/L。
实例2:浸出
某企业高酸浸出铅银渣,含Pb
15.54%(干基),取样588g,加入5L含Ca2+ 14g/L,NaCl
260g/L浸出液,80℃下搅拌浸出1.5小时,浸出渣过滤烘干重451g,含Pb 1.36%.浸出液+洗水共5.32L,含Pb
16.0g/L,Pb浸出率93.15%。
实例3:浸出
废铅酸蓄电池铅膏浸出:某铅冶炼厂蓄电池回收车间,经拆解后的铅泥,含铅76.04%(干基),取铅膏100g,浸出液2L,浸出液成分NaCl
320g/L,HCl
20g/L,浸出温度85℃,浸出时间2小时,浸出终点pH2.0,加入Ca(OH)2调整pH至4.5.过滤后滤液+洗水2.1L,含Pb
34.8g/L,铅浸出率96%。
实例4:浸出
某企业挥发窑氧化锌,经硫酸浸出锌和铟后的铅渣,含Pb21.2%、
Zn7.14%,取渣100g,浸出温度18℃,液固比10:1,浸出时间5小时。浸出后过滤、洗渣,按不同配比NaCl和CaCl2浸出液进行浸出。
浸出结果如下:
序号 | 浸出液 | 浸出后液+洗水 | 浸出率 | ||||
NaCl(g/L) | CaCl2(g/L) | 体积(L) | Pb(g/L) | Zn(g/L) | Pb | Zn | |
1 | 200 | 25 | 1.10 | 3.55 | 2.42 | 18.43 | 37.3 |
2 | 250 | 25 | 1.30 | 5.54 | 1.94 | 34.03 | 35.3 |
3 | 300 | 30 | 1.27 | 10.12 | 2.09 | 60.7 | 36.5 |
4 | 360 | 25 | 1.25 | 14.62 | 2.05 | 86.5 | 35.9 |
5 | 350 | 50 | 1.19 | 17.59 | 2.86 | 98.7 | 47.7 |
在18℃的条件下,NaCl和CaCl2量升高,PbSO4的浸出率提高,在序号5条件下,浸出率达到比较理想。
实例5:净化除杂
PbCl2氯化浸出液,成分如下:
Pb (g/L) | Zn (g/L) | Fe (mg/L) | Cu (mg/L) | As (mg/L) | Sb (mg/L) | Bi (mg/L) | Sn (mg/L) | Ca (g/L) |
21.86 | 7.74 | 10 | 80.79 | 1.0 | 2.05 | 1.79 | 6.37 | 5.71 |
取溶液5L,在50℃条件下,加入锌粉2.5g,搅拌反应,锌粉加入后,很快置换出海绵铅,再继续搅拌,海绵铅在搅拌下形成粉状,再置换杂质,反应1小时后过滤,溶液成分如下:
Pb (g/L) | Zn (g/L) | Fe (mg/L) | Cu (mg/L) | As (mg/L) | Sb (mg/L) | Bi (mg/L) | Sn (mg/L) | Ca (g/L) |
21.50 | 8.01 | 10 | 1.2 | 0.1 | 0.1 | 0.2 | 0.2 | 5.7 |
实例6:铅置换
置换前液成分:
g/L | Pb | Zn | Ca | pH |
22.97 | 6.66 | 3.5 | 5.0 |
取溶液5L,放入容器内,温度50℃,放入锌片200g,快速搅动溶液,5分钟后溶液含Pb降至0.02g/L。取出锌片和海绵铅,残余锌片称重163g,锌片消耗47g,计算置换铅量114.75g。锌片消耗0.3224g
Zn/g Pb,为理论消耗量1.02倍。
置换后液成分:
g/L | Pb | Zn | pH |
0.02 | 14.1 | 5.0 |
实例7:铅置换
PbCl2的氯盐溶液成分:
g/L | Pb | Zn | Ca |
12.72 | 3.94 | 4.30 |
在容器内量入浸出液,开动搅拌加入1000g锌片,溶液连续进、连续出,待锌片消失,置换后铅成粒状,取出后压团、称重,取样分析。团重3210g,含Pb
95.71%,水分4.5%。置换铅量3210×95.71%×95.5%=2934g,置换Pb耗Zn量为0. 3408 g Zn/g
Pb,为理论量108%。
压团用320g NaOH覆盖,在500℃下熔化,得铅锭。铅锭含Pb
99.8%,杂质成分:Zn0.0069%、Fe0.0017%、
Bi0.002%、Cu0.074%、Sb0.0078%、Ag252g/t、As<0.01、Sn<0.01,杂质总和0.1196%。
实例8:置换后液萃取回收锌
萃取剂用工业P204,以260#煤油为稀释剂,按体积配比20%浓度,理论饱和容量分别为Zn2+17.8g/L、Ca2+10.9g/L。
有机相皂化:在皂化液中加入CaCl2,溶液含Ca2+9.83g/L,用20%
P204溶液按1:1相比与皂化液混合4分钟分相,皂化后液含Ca2+0.9g/L,pH4.5;有机相含Ca2+8.9g/L,用于锌萃取,皂化后液加Ca(OH)2将Ca2+补充至10g/L左右,循环使用。
锌萃取:
萃取原液 g/L: | Zn | Ca | Cl- |
14.85 | 4.5 | 188 | |
皂化有机相 g/L: | - | 9.16 | - |
萃取按O/A=1:1.5进行,四级逆流,混合时间3分钟,分相时间1分钟。萃取达到平衡后,取有机相和水相分析,有机相含Zn14.66g/L,含Ca
0.052g/L。萃余液含Ca 10.5g/L、Zn4.2g/L,返回PbSO4渣的浸出。
反萃锌:萃取锌后有机相用含锌的硫酸溶液进行反萃,反萃液锌、酸含量与锌电解酸液相同。为考察反萃液中的Ca和Cl-状况,反萃液用试剂、蒸馏水配制。反萃前后溶液成分:
Zn(g/L) | H2SO4 | Ca(mg/L) | Cl-(mg/L) | |
反萃前液 | 45 | 170 | - | - |
反萃后液 | 92 | 98 | 10.6 | 34.7 |
反萃后液成分符合锌电解要求。
反萃后液经过气浮—活性炭吸附后,可以进行锌电解。锌电解按本企业锌萃取—电解锌工艺产出锌片。
实例9:铅置换后液中和沉锌及Ca(OH)2溶解
中和沉锌为锌回收的第二方案。
铅置换后液成分g/L: | Zn | Pb | Ca |
14.1 | 0.02 | 3.2 |
取置换后液5L,加入粉状Ca(OH)2,在搅拌槽内进行,中和至pH7.0,产出Zn(OH)2
119.1g ,含Zn 59.2%,沉锌后液含Zn
5mg/L。含Ca2+12.3g/L,沉锌后液可返回PbSO4浸出。
Zn(OH)2溶解:为适应后续锌回收需求,采用本公司锌萃取生产的锌萃余液进行溶解。
萃余液成分g/L: | Zn | H2SO4 |
7.5 | 21 |
溶解后液: Zn 21.5g/L、pH5.0
锌溶解率>99%
溶解后液符合锌萃取要求,采用锌萃取—电解锌工艺生产锌片。
Claims (4)
- 一种硫酸铅湿法炼铅工艺,其特征是:以CaCl2 + NaCl为浸出溶剂,对含有硫酸铅的湿法炼锌的高浸渣、含铅氧化锌浸出锌后的铅渣或铅酸蓄电池拆解后的铅泥浸出,浸后液用锌置换铅,置换后液回收锌,回收的锌返回置换铅工序,回收锌后的含Cl-液循环到硫酸铅氯盐浸出工序反复循环使用,其中,①Ca2+摩尔比≥物料中SO4 2-摩尔比;②NaCl和CaCl2的浸出液配比应根据操作中实时监测数据调整,保证浸出后溶液含Ca2+为3~4g/L、含Pb2+ 为10~30g/L,浸出液中Cl-浓度≥150g/L;③锌回收有两种方案,即a.置换后液采用P204为萃取剂,经过萃取剂皂化→萃取锌→洗杂→反萃→电解锌一系列工序,b. 用Ca (OH)2中和沉锌,再经过Zn(OH)2溶解,锌萃取,反萃和电解锌四个工序。
- 按权利要求1所述的硫酸铅湿法炼铅工艺,其特征是:浸出液温度控制在18℃~90℃,浸出液的pH=4~5;浸后液应采用海绵铅净化,除去溶液中的Ag、Cu、Bi微量杂质;置换所用的锌为金属锌片、锌粉、锌粒、锌球,以电解锌析出的锌片最佳,置换温度为18℃~90℃,置换锌用量为理论用量的105~110%。
- 按权利要求1所述的硫酸铅湿法炼铅工艺,其特征是锌回收第一种方案a的具体操作为:①萃取剂皂化:用Ca(OH)2对P204进行皂化,P204与260#溶剂油的体积比为20~30%,用配制好的P204与皂化剂混合均匀1~2分钟,静置沉清分相,皂化后液加入Ca(OH)2后循环使用;②萃取锌:皂化后的P204与置换后液混合,澄清分相,混合萃取时间1~5分钟,采用3~5级逆流萃取,萃取温度20~40℃,萃取后有机相中夹带的残余萃余液用水洗涤除去;③反萃锌:萃锌后有机相用锌电解废液进行反萃,锌电解废液含H2SO4 150~180g/L,含Zn40~50g/L,反萃相比按照反萃后含Zn 90~120g/L进行控制,反萃按1~2级逆流进行,以有机相含Zn<0.5g/L为准,反萃混合时间0.5~10分钟,反萃温度20~40℃,反萃后的有机相返回皂化工序循环使用,反萃液中夹带的残余萃取剂采用气浮——活性炭吸附;④电解锌:反萃得到的ZnSO4液进行电解处理,再生的H2SO4返回反萃锌循环使用,析出的锌片用于置换铅工序,多余回收的锌片铸锭后作为锌产品。
- 按权利要求1所述的硫酸铅湿法炼铅工艺,其特征是锌回收第二种方案b的具体操作为:①中和沉锌:用Ca (OH)2中和沉锌,使pH升高至7~9,然后液固分离,沉锌后液返回PbSO4的浸出工序,循环使用,分离出Zn(OH)2沉淀物,再用水洗除夹带的母液;Zn(OH)2可以制成ZnO产品,也可以按下述过程生产电锌。②Zn(OH)2溶解:在搅拌槽内用稀硫酸溶解Zn(OH)2,控制浸出液含Zn 20~30g/L,溶解锌后溶液过滤后进行锌萃取;③萃取锌:以260#溶剂煤油为稀释剂与P204混合,配制成P204体积含量为20~40%的有机相,用有机相与溶解锌后溶液搅拌混合,有机相与水相比O/A = 1/2~2/1,萃取混合时间0.5~5分钟,萃取温度20~30℃,进行3~4级逆流萃取,使锌转入有机相,萃余液返回Zn(OH)2溶解,循环使用;④反萃锌:与锌回收第一种方案对应相同;⑤电解锌:与锌回收第一种方案对应相同。
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