CN102206750A - Method for recovering lead from lead-containing material by matching leaching-electrowinning method - Google Patents
Method for recovering lead from lead-containing material by matching leaching-electrowinning method Download PDFInfo
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- CN102206750A CN102206750A CN2011101142133A CN201110114213A CN102206750A CN 102206750 A CN102206750 A CN 102206750A CN 2011101142133 A CN2011101142133 A CN 2011101142133A CN 201110114213 A CN201110114213 A CN 201110114213A CN 102206750 A CN102206750 A CN 102206750A
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- lead
- leaching
- containing material
- plumbous
- electrodeposition
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- 238000000034 method Methods 0.000 title claims abstract description 62
- 239000000463 material Substances 0.000 title claims abstract description 26
- 238000005363 electrowinning Methods 0.000 title claims abstract description 5
- 238000002386 leaching Methods 0.000 claims abstract description 46
- KCXVZYZYPLLWCC-UHFFFAOYSA-N EDTA Chemical compound OC(=O)CN(CC(O)=O)CCN(CC(O)=O)CC(O)=O KCXVZYZYPLLWCC-UHFFFAOYSA-N 0.000 claims abstract description 38
- 239000011701 zinc Substances 0.000 claims abstract description 37
- XLOMVQKBTHCTTD-UHFFFAOYSA-N Zinc monoxide Chemical group [Zn]=O XLOMVQKBTHCTTD-UHFFFAOYSA-N 0.000 claims abstract description 29
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 claims abstract description 27
- 229910052725 zinc Inorganic materials 0.000 claims abstract description 27
- 238000004070 electrodeposition Methods 0.000 claims abstract description 23
- 239000002893 slag Substances 0.000 claims abstract description 19
- 239000003795 chemical substances by application Substances 0.000 claims abstract description 12
- 239000007788 liquid Substances 0.000 claims abstract description 12
- 239000002699 waste material Substances 0.000 claims abstract description 12
- 239000003513 alkali Substances 0.000 claims abstract description 10
- YEXPOXQUZXUXJW-UHFFFAOYSA-N oxolead Chemical compound [Pb]=O YEXPOXQUZXUXJW-UHFFFAOYSA-N 0.000 claims abstract description 9
- WABPQHHGFIMREM-UHFFFAOYSA-N lead(0) Chemical compound [Pb] WABPQHHGFIMREM-UHFFFAOYSA-N 0.000 claims abstract description 8
- PIJPYDMVFNTHIP-UHFFFAOYSA-L lead sulfate Chemical compound [PbH4+2].[O-]S([O-])(=O)=O PIJPYDMVFNTHIP-UHFFFAOYSA-L 0.000 claims abstract description 5
- 238000000926 separation method Methods 0.000 claims abstract description 5
- HWSZZLVAJGOAAY-UHFFFAOYSA-L lead(II) chloride Chemical compound Cl[Pb]Cl HWSZZLVAJGOAAY-UHFFFAOYSA-L 0.000 claims abstract description 4
- 229910052751 metal Inorganic materials 0.000 claims abstract description 4
- 239000002184 metal Substances 0.000 claims abstract description 4
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 21
- 239000000292 calcium oxide Substances 0.000 claims description 10
- ODINCKMPIJJUCX-UHFFFAOYSA-N calcium oxide Inorganic materials [Ca]=O ODINCKMPIJJUCX-UHFFFAOYSA-N 0.000 claims description 10
- 235000019504 cigarettes Nutrition 0.000 claims description 10
- 238000002156 mixing Methods 0.000 claims description 10
- 238000000498 ball milling Methods 0.000 claims description 7
- 238000003723 Smelting Methods 0.000 claims description 6
- BRPQOXSCLDDYGP-UHFFFAOYSA-N calcium oxide Chemical compound [O-2].[Ca+2] BRPQOXSCLDDYGP-UHFFFAOYSA-N 0.000 claims description 6
- RTAQQCXQSZGOHL-UHFFFAOYSA-N Titanium Chemical compound [Ti] RTAQQCXQSZGOHL-UHFFFAOYSA-N 0.000 claims description 5
- 229940009662 edetate Drugs 0.000 claims description 5
- 239000000843 powder Substances 0.000 claims description 5
- 229910001220 stainless steel Inorganic materials 0.000 claims description 5
- 239000010935 stainless steel Substances 0.000 claims description 5
- 239000010936 titanium Substances 0.000 claims description 5
- 229910052719 titanium Inorganic materials 0.000 claims description 5
- BEGBSFPALGFMJI-UHFFFAOYSA-N ethene;sodium Chemical group [Na].C=C BEGBSFPALGFMJI-UHFFFAOYSA-N 0.000 claims description 4
- 239000000126 substance Substances 0.000 claims description 4
- 238000013019 agitation Methods 0.000 claims description 3
- 238000004519 manufacturing process Methods 0.000 claims description 3
- 239000000428 dust Substances 0.000 claims description 2
- 150000004965 peroxy acids Chemical class 0.000 claims description 2
- -1 plumbum ion Chemical class 0.000 claims description 2
- 230000001376 precipitating effect Effects 0.000 claims description 2
- 239000007787 solid Substances 0.000 claims description 2
- 239000011787 zinc oxide Substances 0.000 abstract description 14
- 238000005265 energy consumption Methods 0.000 abstract description 8
- 238000001914 filtration Methods 0.000 abstract description 3
- 238000011282 treatment Methods 0.000 abstract description 3
- 238000002360 preparation method Methods 0.000 abstract description 2
- 239000010865 sewage Substances 0.000 abstract description 2
- 229940071106 ethylenediaminetetraacetate Drugs 0.000 abstract 1
- 229910000464 lead oxide Inorganic materials 0.000 abstract 1
- 239000004071 soot Substances 0.000 abstract 1
- 239000000243 solution Substances 0.000 description 17
- 238000006243 chemical reaction Methods 0.000 description 8
- HTUMBQDCCIXGCV-UHFFFAOYSA-N lead oxide Chemical compound [O-2].[Pb+2] HTUMBQDCCIXGCV-UHFFFAOYSA-N 0.000 description 6
- 239000002253 acid Substances 0.000 description 4
- 238000005868 electrolysis reaction Methods 0.000 description 4
- QTBSBXVTEAMEQO-UHFFFAOYSA-N Acetic acid Chemical compound CC(O)=O QTBSBXVTEAMEQO-UHFFFAOYSA-N 0.000 description 3
- 239000000460 chlorine Substances 0.000 description 3
- UQLDLKMNUJERMK-UHFFFAOYSA-L di(octadecanoyloxy)lead Chemical compound [Pb+2].CCCCCCCCCCCCCCCCCC([O-])=O.CCCCCCCCCCCCCCCCCC([O-])=O UQLDLKMNUJERMK-UHFFFAOYSA-L 0.000 description 3
- 229960001484 edetic acid Drugs 0.000 description 3
- 238000001556 precipitation Methods 0.000 description 3
- 239000000047 product Substances 0.000 description 3
- 150000003839 salts Chemical class 0.000 description 3
- 238000005303 weighing Methods 0.000 description 3
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 2
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 description 2
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 description 2
- OSGAYBCDTDRGGQ-UHFFFAOYSA-L calcium sulfate Chemical compound [Ca+2].[O-]S([O-])(=O)=O OSGAYBCDTDRGGQ-UHFFFAOYSA-L 0.000 description 2
- 230000007613 environmental effect Effects 0.000 description 2
- 238000003912 environmental pollution Methods 0.000 description 2
- JEHCHYAKAXDFKV-UHFFFAOYSA-J lead tetraacetate Chemical compound CC(=O)O[Pb](OC(C)=O)(OC(C)=O)OC(C)=O JEHCHYAKAXDFKV-UHFFFAOYSA-J 0.000 description 2
- 230000035484 reaction time Effects 0.000 description 2
- 238000011084 recovery Methods 0.000 description 2
- 229910052709 silver Inorganic materials 0.000 description 2
- 239000004332 silver Substances 0.000 description 2
- MFEVGQHCNVXMER-UHFFFAOYSA-L 1,3,2$l^{2}-dioxaplumbetan-4-one Chemical compound [Pb+2].[O-]C([O-])=O MFEVGQHCNVXMER-UHFFFAOYSA-L 0.000 description 1
- OKTJSMMVPCPJKN-UHFFFAOYSA-N Carbon Chemical compound [C] OKTJSMMVPCPJKN-UHFFFAOYSA-N 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
- 239000003109 Disodium ethylene diamine tetraacetate Substances 0.000 description 1
- CWYNVVGOOAEACU-UHFFFAOYSA-N Fe2+ Chemical compound [Fe+2] CWYNVVGOOAEACU-UHFFFAOYSA-N 0.000 description 1
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 1
- 229910000003 Lead carbonate Inorganic materials 0.000 description 1
- 241001365789 Oenanthe crocata Species 0.000 description 1
- 229910004298 SiO 2 Inorganic materials 0.000 description 1
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 1
- 239000005864 Sulphur Substances 0.000 description 1
- 239000000654 additive Substances 0.000 description 1
- 230000000996 additive effect Effects 0.000 description 1
- 239000012670 alkaline solution Substances 0.000 description 1
- 239000006227 byproduct Substances 0.000 description 1
- 239000011575 calcium Substances 0.000 description 1
- 229910052799 carbon Inorganic materials 0.000 description 1
- 229910052801 chlorine Inorganic materials 0.000 description 1
- 239000012141 concentrate Substances 0.000 description 1
- 238000005137 deposition process Methods 0.000 description 1
- 235000019301 disodium ethylene diamine tetraacetate Nutrition 0.000 description 1
- 238000001035 drying Methods 0.000 description 1
- 230000000694 effects Effects 0.000 description 1
- 238000003411 electrode reaction Methods 0.000 description 1
- 238000005516 engineering process Methods 0.000 description 1
- 238000000605 extraction Methods 0.000 description 1
- 239000000706 filtrate Substances 0.000 description 1
- 239000012467 final product Substances 0.000 description 1
- 239000003517 fume Substances 0.000 description 1
- 238000009854 hydrometallurgy Methods 0.000 description 1
- 239000012535 impurity Substances 0.000 description 1
- 239000013067 intermediate product Substances 0.000 description 1
- 229910052742 iron Inorganic materials 0.000 description 1
- SMBGWMJTOOLQHN-UHFFFAOYSA-N lead;sulfuric acid Chemical compound [Pb].OS(O)(=O)=O SMBGWMJTOOLQHN-UHFFFAOYSA-N 0.000 description 1
- 238000010310 metallurgical process Methods 0.000 description 1
- 238000005554 pickling Methods 0.000 description 1
- 239000004033 plastic Substances 0.000 description 1
- 229920003023 plastic Polymers 0.000 description 1
- 230000001698 pyrogenic effect Effects 0.000 description 1
- 238000007086 side reaction Methods 0.000 description 1
- 239000010802 sludge Substances 0.000 description 1
- RYYKJJJTJZKILX-UHFFFAOYSA-M sodium octadecanoate Chemical compound [Na+].CCCCCCCCCCCCCCCCCC([O-])=O RYYKJJJTJZKILX-UHFFFAOYSA-M 0.000 description 1
- 239000007790 solid phase Substances 0.000 description 1
- 238000007614 solvation Methods 0.000 description 1
- 238000005406 washing Methods 0.000 description 1
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 1
- NWONKYPBYAMBJT-UHFFFAOYSA-L zinc sulfate Chemical compound [Zn+2].[O-]S([O-])(=O)=O NWONKYPBYAMBJT-UHFFFAOYSA-L 0.000 description 1
- 229960001763 zinc sulfate Drugs 0.000 description 1
- 229910000368 zinc sulfate Inorganic materials 0.000 description 1
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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
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- Electrolytic Production Of Metals (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
本发明公开了一种从含铅物料中配合浸出-电积法回收铅的方法，采用0.05-0.5M的乙二胺四乙酸盐加0.05-1.0M的碱为浸出剂，在常温-95℃的温度下，从含铅物料中选择性浸出铅，过滤分离后得到的含铅溶液，再采用电积法从溶液中析出金属铅粉，电积废液经配液后返回浸出。对于次氧化锌烟灰，通过浸出铅可以提高次氧化锌中锌的品位与后续浸出的浸出率，便于后序处理；对于含氧化铅、硫酸铅或氯化铅的低品位铅渣，浸出铅后，渣堆存过程中沥出的污水将不含或少含铅，对环境不产生影响，另一方面，变废为宝，可以回收铅渣中的铅。该方法具有流程简单、易于操作、低能耗等优点，能广泛处理含铅次氧化锌物料及各种低品位铅渣。The invention discloses a method for recovering lead from lead-containing materials combined with the leaching-electrodeposition method, using 0.05-0.5M ethylenediamine tetraacetate plus 0.05-1.0M alkali as the leaching agent, and at room temperature -95 At a temperature of ℃, lead is selectively leached from lead-containing materials, and the lead-containing solution obtained after filtration and separation is used to precipitate metal lead powder from the solution by electrowinning, and the electrowinning waste solution is returned to leaching after liquid preparation. For secondary zinc oxide soot, the grade of zinc in secondary zinc oxide and the leaching rate of subsequent leaching can be improved by leaching lead, which is convenient for subsequent treatment; for low-grade lead slag containing lead oxide, lead sulfate or lead chloride, after leaching lead , The sewage leached during the slag storage process will contain no or less lead, and will not affect the environment. On the other hand, the lead in the lead slag can be recovered by turning waste into treasure. The method has the advantages of simple process, easy operation, low energy consumption, etc., and can widely process lead-containing sub-zinc oxide materials and various low-grade lead slags.
Description
Technical field
A kind of alkaline solution of edetate that adopts disclosed by the invention cooperates the lead, the electrodeposition that leach in the lead-containing material to reclaim plumbous novel process, belongs to secondary resource and reclaims and non-ferrous metal hydrometallurgy field.
Background technology
Secondary zinc oxide cigarette ash is a kind of intermediate product plumbous, the zinc metallurgical process, is to contain zinc, lead fume dirt by being deposited on of producing in the lead blast furnace slag blast furnace fuming process in the flue.Behind general at present process multiple hearth furnace earlier or pickling defluorinate, the chlorine, be used for the Production By Sulfuric Acid Process metallic zinc, lead enters in the acid leaching residue, obtain the sulfuric acid lead skim of leaded 10-30%, low-grade lead skim that obtains and lead ore concentrate batching enter the plumbous smelting of pyrogenic process, and the rate of recovery of this method zinc is lower, and plumbous zinc circulates in the centre, strengthen logistics capacity, and energy consumption height, environmental pollution are more serious.
At present, for hydrometallurgic recovery lead from contain lead skim, Chen Wei equality adopts acid solution (using 10% HCl) to leach plumbous (Chen Weiping under leaching under high temperature, the high pressure and ultrasonic wave condition, Gong Jiansen. the China YouSe Acta Metallurgica Sinica, 1996,6 (3): 43-46), but this method leaching agent corrodibility is big, and is higher to equipment requirements.People such as Xi Changsheng propose with lead containing sludge prepare lead stearate reclaim plumbous (Xi Changsheng, Chen Guosheng, Peng Cuihong etc. Guangdong chemical industry, 2001,6 (11): 44-46), utilize the difference of different indissoluble lead salt ionic product concentration to realize the lead salt conversion.At first utilize the reaction of volatile salt and lead sulfate, generate the lead carbonate precipitation,, obtain plumbic acetate solution with dilute acetic acid dissolved carbon lead plumbate; Filtrate is added sodium stearate solution, generate the lead stearate precipitation, washing, precipitation separation, drying gets final product to such an extent that be widely used in the lead stearate product of plastics additive, also can make plumbic acetate as required simultaneously.This method is adapted to the lower lead skim of content, have to invest characteristics little, high efficiency, but treating processes is too complicated, not easy to operate.People such as Guo Cuixiang propose with NaOH handle leaded secondary zinc oxide slag (Guo Cuixiang, Zhao Youcai etc. chemical industry environmental protection, 2008,28 (1): 77-80), the plumbous zinc of leached mud the inside utilizes potential difference to select the plumbous zinc of electrodeposition to reach the isolating purpose of plumbous zinc then simultaneously.This method has environmental protection, and is efficient, less energy-consumption, and characteristics such as flow process is simple, but the NaOH concentration requirement is very high, filtration difficulty, and plumbous zinc enters in the leach liquor together, and the lead that contains 2%-3% behind the electrodeposition all the time also need further carry out purifying treatment.And the very low lead skim of content contains Pb as waste dump in the drop water outlet
2+, this causes very big harm to environment and human health.
Summary of the invention
The purpose of this invention is to provide a kind of leaching-electrodeposition method that cooperates and reclaim plumbous method from lead-containing material, this method can improve the grade of zinc in the secondary zinc oxide and the leaching yield of follow-up leaching by leaching lead, is convenient to postorder and handles; And be fit to handle the low-grade lead skim that contains plumbous oxide, lead sulfate or lead chloride, leach lead after, slag muck is deposited that drop goes out in the process sewage will not contain or is few leaded, and environment is not exerted an influence, and can reclaim the lead in the lead skim simultaneously, turns waste into wealth.
The objective of the invention is to realize in the following manner.
A kind of leaching-electrodeposition method that cooperates from lead-containing material reclaims plumbous method: adopting the mixing solutions of edetate and alkali is leaching agent, selectivity leaches plumbous from lead-containing material, what obtain after the filtering separation contains lead solution, adopts electrodeposition method precipitating metal lead powder from solution again.
Described lead-containing material comprises various one or more materials that contain in plumbous oxide, lead sulfate, the lead chloride.As: the acidleach that derives from plumbum smelting slag cigarette ash heat zinc oxide fume dust, the generation of production zinc chemical product is slagged tap or zinc smelting high temperature peracid leached mud.
The mixing solutions of the edetate of described leaching agent employing 0.05-0.5M and the alkali of 0.05-1.0M; Alkali comprises: sodium hydroxide or calcium oxide.
Described leaching agent can be the mixing solutions of mixing solutions, Ca-EDTA and the calcium oxide of sodium ethylene diamine tetracetate and sodium hydroxide, or the mixing solutions of sodium ethylene diamine tetracetate and calcium oxide.
The liquid-solid ratio of described leaching agent and lead-containing material is 1L: 100-300g.
Described leaching process adopts agitation leach or ball milling to leach.Especially adopt CaEDTA to leach PbSO
4In the process, because the CaSO that generates
4Be coated on plumbous surface, influence leaching yield, need to adopt the method for the ball milling while leaching, temperature is a normal temperature, and ball is 1kg: 5-50L with the ratio of slip.
Described leaching process temperature is 0-95 ℃, and the time is 0.5-3h.
Adopting stainless steel or titanium plate in the described electrowinning lead powder process is negative electrode, the steady anode of type, and at 20-70 ℃, current density 50-500A/m
2Condition under, electrodeposition plumbum ion concentration 1-2g/L to the solution produces the lead bullion powder on the negative electrode, collect the back founding and become lead bullion.
Waste liquid behind the described electrodeposition returns leaching after dosing.
Ultimate principle of the present invention is:
PbO+EDTA
2-+H
2O＝Pb(EDTA)+2OH
- (1)
ZnO+EDTA
2-+H
2O＝Zn(EDTA)+2OH
- (2)
Zn(EDTA)+PbO+H
2O＝Pb(EDTA)+Zn(OH)
2 (3)
PbSO
4+EDTA
2-＝Pb(EDTA)+SO
4 2- (4)
PbCl
2+EDTA
2-＝Pb(EDTA)+2Cl
- (5)
PbSO
4+EDTA
2-+CaO+H
2O＝Pb(EDTA)+CaSO
4+2OH
- (6)
Because of Pb
2+-EDTA
2-Title complex compares Zn
2+-EDTA
2-The complexes stability height is so there is being Pb
2+Can suppress Zn when existing
2+Leaching.Leach in the secondary zinc oxide cigarette ash process, even because zinc content height in the cigarette ash, zinc optimum solvation, along with the prolongation of extraction time, also can be slowly with solid phase in plumbous exchange enter in the slag.But when leaching lead-containing material for containing PbSO
4During material, the SO during plumbous the dissolving in the material
4 2-Also enter solution, influence the balance of reaction formula (4), thus the plumbous leaching yield of influence; Therefore will add a certain amount of alkali improves Pb
2+Leaching.
Temperature one is K regularly
pConstant, add calcium oxide and generate calcium sulfate, reduce sulfate radical, can improve PbEDTA certainly.For PbCl
2Leach this and will explain that ethylenediamine tetraacetic acid (EDTA) is a weak acid from the characteristic of ethylenediamine tetraacetic acid (EDTA), pH is lower than 5 just can not be dissociated into EDTA fully
2-, just do not form title complex.Leaching PbCl
2In the material process, the muriate of some impurity (as iron(ic) chloride) can consume alkali, and system pH is reduced, and therefore need add alkali improves the pH value; Sulphur in the secondary zinc oxide cigarette ash is to exist with sulfate radical in fact, so also will add alkali, is less; In a word, ethylenediamine tetraacetic acid (EDTA) disalt of the present invention is used for cooperating plumbous; The effect of sodium hydroxide or CaO is to adjust the SO in the leaching process
4 2-, Cl
-To Pb
2+Influence.
Because equilibrium potential plumbous, zinc differs bigger, therefore can adopt the method selective electrolysis lead of direct electrodeposition.Electrode reaction can be expressed as.
Anode: 2H
2O-4e
-=4H
++ O
2(7)
Negative electrode: Pb (EDTA)+2e
-=Pb+EDTA
2-(8)
Negative electrode side reaction: Zn (EDTA)+2e
-=Zn+EDTA
2-(9)
Cu(EDTA)+2e
-＝Cu+EDTA
2- (10)
Because general leaching agent EDTA
2-Therefore content with respect to lead is excessive, has part zinc to enter in the solution, because the influence of the kinetic factor of mass transfer velocity, a little zinc also electrodeposition enters in the metallic lead in electrolytic deposition process.
The whole plumbous electrolytic process current efficiency of the present invention is 82%-98%, and energy consumption is a 600-1000kWh/t lead.
The invention has the advantages that:
(1) for adopting EDTA
2-Selectivity leaches plumbously from leaded secondary zinc oxide cigarette ash, has plumbously obtained good the separation with zinc, no matter is the further electrodeposition for follow-up lead, or the further processing of zinc, has all established good basis, has improved the efficient of subsequent operations greatly;
(2) the leaching agent preparation of using is simple, and concentration requirement is not high, and is therefore, not high to equipment requirements yet;
(3) whole process flow is simple, easy handling; Energy consumption is low, and lead content reduces greatly in the slag, reduces environmental pollution;
(4) can leaded secondary zinc oxide material of extensive treatments and the various low-grade lead skim that contains.
Description of drawings
Fig. 1 is the concrete process flow sheet of the present invention.
Embodiment
Be intended to further specify the present invention below in conjunction with embodiment, and unrestricted the present invention.
Embodiment 1:
Certain factory plumbous smelting by-product secondary zinc oxide cigarette ash, material component (%): Zn 52.43, and Pb 12.44, and Cu 0.04, and In 0.0199, and S 3.2, and Cl 0.06, and F 0.084, and As 0.32, and Sb 0.047.
Disodium ethylene diamine tetraacetate 0.15mol/L, the solution of NaOH 0.2mol/L, take by weighing secondary zinc oxide cigarette ash by solid-to-liquid ratio 200g/L, add in the reactor, control reaction temperature is 90 ℃, and agitation leach 2.0h makes the lead in the secondary zinc oxide cigarette ash enter in the leach liquor with the form of Pb (EDTA) title complex, slag contains Pb 1.06%, and pH of leaching solution is 10.Leaching yield plumbous in the leaching process is 93.65%, and zinc leaching rate only is 4.24%.
The leaded 22.04g/L of leach liquor, adopt stainless steel as cloudy, be coated with PbO
2The titanium plate is an anode, and at 20-30 ℃, current density is 100A/m
2Under carry out electrodeposition, voltage is 1.7-2.2V electrolysis 10h, the leaded 1.8g/L of waste liquid obtains the lead bullion powder of leaded 88.4%, zinc 4.6% and water 6.8% on negative electrode, whole plumbous electrolytic process current efficiency is 90%, energy consumption is a 810kWh/t lead.Return after the dosing of electrodeposition waste liquid next time and leach.
Embodiment 2:
Certain zinc sulfate factory produces the acidleach lead skim, and material component (%): Zn 3.89, Pb18.16, and In 0.025, SO
4 2-14.2, Fe0.56, Cu 0.05, and Sb 1.47, and Bi 0.85, and Ca 1.20.
CaEDTA 0.3mol/L, the solution of CaO 1.0mol/L is that 150g/L takes by weighing and contains lead skim by solid-to-liquid ratio, adds together in the ball-milling reaction device, and control reaction temperature is a normal temperature, and ball is 1kg: 10L with the ratio of slip, and ball-milling reaction time 0.5h makes PbSO
4Lead in the slag enters in the leach liquor SO with the form of PbEDTA title complex
4 2-With CaSO
4The form of slag remains in the slag, and leached mud contains Pb 0.64%, and leaching yield plumbous in the leaching process is 96.65%.
The leaded 27.08g/L of leach liquor, adopt stainless steel as cloudy, be coated with PbO
2The titanium plate is an anode, and at 40 ℃, current density is 400A/m
2Under carry out electrodeposition, voltage is 2.0-2.5V electrolysis 12h, the leaded 1.2g/L of waste liquid obtains leaded 92.0%, as to contain zinc 0.6% lead bullion powder on negative electrode, whole plumbous electrolytic process current efficiency is 92%, energy consumption is a 630kWh/t lead.Return after the dosing of electrodeposition waste liquid next time and leach.
Embodiment 3:
Certain wet method zinc factory produces high acid leaching residue, and material component (%) is: Zn 8.21, and Pb 7.76, and Ag 0.162, and Ge 0.015, Fe13.76, SO
4 2-4.2, SiO
225.34.
CaEDTA 0.4mol/L, the solution of CaO 0.5mol/L is that 300g/L takes by weighing and contains lead skim by solid-to-liquid ratio, adds together in the ball-milling reaction device, and control reaction temperature is a normal temperature, and ball is 1kg: 30L with the ratio of slip, and ball-milling reaction time 1.0h makes PbSO
4Lead in the slag enters in the leach liquor SO with the form of PbEDTA title complex
4 2-With CaSO
4The form of slag remains in the slag, and leached mud contains Pb 0.48%, and leaching yield plumbous in the leaching process is 93.52%, and the leaching yield of silver is 85%.
The leaded 21.24g/L of leach liquor, adopt stainless steel as cloudy, be coated with IrO
2The titanium plate is an anode, and at 45 ℃, current density is 200A/m
2Under carry out electrodeposition, voltage is 1.7-2.2V electrolysis 24h, the leaded 1.5g/L of waste liquid obtains the lead bullion powder of leaded 84.0%, silver 1.8%, zinc 5.6% on negative electrode, whole plumbous electrolytic process current efficiency is 82%, energy consumption is a 980kWh/t lead.Return after the dosing of electrodeposition waste liquid next time and leach.
Claims (10)
1. one kind cooperates leaching-electrodeposition method to reclaim plumbous method from lead-containing material, it is characterized in that adopting the mixing solutions of edetate and alkali is leaching agent, selectivity leaches plumbous from lead-containing material, what obtain after the separation contains lead solution, adopts electrodeposition method precipitating metal lead powder from solution again.
2. method according to claim 1 is characterized in that, described lead-containing material comprises various one or more materials that contain in plumbous oxide, lead sulfate, the lead chloride.
3. method according to claim 1 and 2 is characterized in that, described lead-containing material derives from the acidleach of plumbum smelting slag cigarette ash heat zinc oxide fume dust, the generation of production zinc chemical product and slags tap or zinc smelting high temperature peracid leached mud.
4. method according to claim 1 is characterized in that, the mixing solutions of the edetate of described leaching agent employing 0.05-0.5M and the alkali of 0.05-1.0M.
5. according to claim 1 or 4 described methods, it is characterized in that, described leaching agent is the mixing solutions of mixing solutions, Ca-EDTA and the calcium oxide of sodium ethylene diamine tetracetate and sodium hydroxide, or the mixing solutions of sodium ethylene diamine tetracetate and calcium oxide.
6. method according to claim 5 is characterized in that, the liquid-solid ratio of described leaching agent and lead-containing material is 1L: 100-300g.
7. method according to claim 1 is characterized in that, described leaching process adopts agitation leach or ball milling to leach.
8. according to claim 1 or 7 described methods, it is characterized in that described leaching process temperature is 0-95 ℃, the time is 0.5-3h.
9. method according to claim 1 is characterized in that, adopting stainless steel or titanium plate in the described electrowinning lead powder process is negative electrode, the steady anode of type, and at 20-70 ℃, current density 50-500A/m
2Condition under, electrodeposition plumbum ion concentration 1-2g/L to the solution produces the lead bullion powder on the negative electrode, collect the back founding and become lead bullion.
10. according to claim 1 or 9 described methods, it is characterized in that the waste liquid behind the described electrodeposition returns leaching after dosing.
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