CN1014722B - New technology of wet lead-smelting - Google Patents
New technology of wet lead-smeltingInfo
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- CN1014722B CN1014722B CN88105063A CN88105063A CN1014722B CN 1014722 B CN1014722 B CN 1014722B CN 88105063 A CN88105063 A CN 88105063A CN 88105063 A CN88105063 A CN 88105063A CN 1014722 B CN1014722 B CN 1014722B
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- Prior art keywords
- lead
- chloride
- ore
- slag
- flotation
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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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Abstract
The present invention relates to a novel wet metallurgy technology for extracting metal lead from sulfurized lead concentrate ores. The technological process is simple, the ore concentrates are converted into lead chloride in a solid phase way and is floated and purified, lead chloride water solution is electrolyzed by an ionic membrane, and thus, metal lead which conforms to the international standard is prepared. The novel technology has no need of further purifying the solution, and thus, the metal lead with high quality can be directly produced from lead concentrate. The operating temperature of the process is low, the device has no need of special corrosion resistance, the investment is saved, the present invention has no poison or three wastes in the process, the technical economical index is advanced, and the concomitant valent metals, copper, silver, zinc, sulfur, etc. in the ore concentrates can be totally recovered. The process processing scale can be adjusted, and the present invention has favorable adaptability.
Description
The present invention is a kind of hydrometallurgical processes that reclaims metallic lead from leaded sulphide ores.
At present, the method for producing metallic lead in the world is pyrogenic process substantially.But along with the continuous progress of various countries' expanding economy, science and technology, to environmental protection requirement strictness and power supply anxiety day by day.This traditional method then exists insoluble problem.Low to atmospheric pollution, energy utilization rate as low-concentration sulfur dioxide, even more serious labor condition is poor, and flying upward of lead powder dirt in the production process and distributing of lead steam cause producer's lead poisoning.So lead poisoning is the chronic illness that domestic and international lead refinery does not all free in varying degrees.
For overcoming above-mentioned drawback, topsoil and saturnine problem have fundamentally been solved.Since the nearly century, constantly the someone seeks a kind ofly by wet processing, does not need the new way of refining, direct production metallic lead.Particularly in recent years, many countries to the wet processing lead glance, carried out extensive studies.The various technical schemes of metallic lead have been proposed to reclaim from principle.Rule are received following several respects: lead sulfide fused salt electrolysis, villaumite leachings-fused salt electrolysis, villaumite leaching-aqueous electrolysis and suppress lead ore concentrate powder anode electrolysis etc.Ha Sen company research as, U.S. is a kind of from higher-grade sulphide ores cryogenic chlorinating roasting, subsequently with the leaching of thermal chlorination sodium solution, crystallisation by cooling separate out lead chloride, fused salt electrolysis gets metallic lead.
In addition, U.S. Reynolds Centre De Recherches Metallurgiques at first discloses concentrate of lead sulfide ore and has leached crystallization, the patent of molten salt electrolysis method.The patent No. is 3929597, this method is the higher-grade lead ore concentrate, with iron trichloride and common salt aqueous solution, 100 ℃ were leached 15 minutes down, crystallisation by cooling lead chloride, fused salt electrolysis obtain metallic lead.In 8 or five years the 15 international ore dressing meeting paper selected works, published relevant French Trappes(Tuo Laipusi) mining and metallurgy institute, with iron trichloride leach concentrate, fluid purification, crystallisation by cooling get lead chloride, again after dissolving, half industrialized test of electrolysis with ion-exchange film, acquisition metallic lead.
In addition, also there are similar patent in Japan and Sweden.The plumbous method of these wet methods refining all is based on a lead glance chlorination and leaches in a word, cooling then, crystallization, separation lead chloride or send electrolytic reduction to produce metallic lead.Because the solubleness of lead chloride when room temperature is low, so all under the condition of high temperature and big liquid-solid ratio, carry out.Make system devices such as leaching, filtration, dense, storage and liquid conveying huge, the standard that the equipment requirements corrosion resistance nature is strong, power consumption is high, amount of electrolyte does not reach corrosion level plumbous (Ag＜0.001%), further refining.The comprehensive problem that reclaims of next also unresolved valuable metal.
The objective of the invention is:
1,, adopt wet type solid phase conversion process to replace the wet type extract technology according to lead compound infusible characteristics.Make the solid phase lead glance in the medium of the aqueous solution, change into the lead chloride of solid phase.Like this, do not need high temperature, do not need big liquid-solid ratio again, and as returning use behind the solution oxide of medium.
2, seek the method for floating of less energy-consumption, separate unreacted sulfide and other impurity, the purification lead chloride.
3, select the structure of suitable ion-exchange membrane electrolyzer, negative electrode is separated out spongy lead from lead chloride solution, realizes the less energy-consumption electrolysis.
4, select rational technical qualification, help the recovery of valuable metals such as silver, copper, zinc sulphur.
Wet method refining splicer skill of the present invention is made up of following each procedure: at first be that the lead ore concentrate solid phase transforms, must transform slag and conversion fluid.The flotation removal of impurities again of conversion slag, flotation tailings are purified lead chlorides, and lead chloride is made catholyte.Conversion fluid is done anolyte after extracting valuable metals such as zinc, silver, copper, sulphur, in having the electrolyzer of ion-exchange membrane, and the electrolytic deposition spongy lead.
One, the solid phase of lead ore concentrate transforms: be earlier that 0.077mm carries out the solid phase conversion with soaking ore deposit agent (iron trichloride, sodium chloride solution) then with the levigate granularity of concentrate of lead sulfide ore: reaction formula is:
Process control liquid-solid ratio=4～8: 1,25～75 ℃ of temperature are soaked the ore deposit agent and are contained (g/l) Fe
+ 340～80, cl200～25, PH＜1,, intense mechanical stirs, transformation time 1～4 hour.After the liquid-solid separation, obtain transforming slag (lead chloride, elementary sulfur and a little unconverted sulphide ores) and transform back liquid (ferrous chloride, sodium-chlor and copper, silver, zinc etc.).Because lead chloride is insoluble compound, solid phase transforms has avoided former high temperature and big liquid-solid ratio leaching process.
It is different with form to look the content that changes silver in the slag, needs will transform the sodium-chlor liquid that slag places 4mol sometimes again, after a small amount of thin lead powder of adding is adjusted the flotation performance of silver, and capable again flotation removal of impurities.
Transform back liquid, the ferric ion in the liquid fully be reduced into ferrous ion with concentrate of lead sulfide ore.Cooling back iron method of replacement reclaims silver, copper and TBP extraction and reclaims zinc.Remove or part remove lead, silver, copper, zinc ferrous chloride solution, be sent to tank room and make anolyte.
Transform and be adjusted to have in the flute profile conical bottom tank of churned mechanically garden and carry out.Groove inner bushing ring epoxy resins or rubber anti-corrosive, and be furnished with corresponding heating installation.
Two, flotation removal of impurities: floatation process control ore pulp (conversion slag) concentration (weight ratio) 10～25%, PH=2～3, normal-temperature operation, collecting agent adopts 2# oil (terpenol) and ethyl xanthate.Through roughly select, scan with selected after, mine tailing-lead chloride quality can reach more than the Pb65%.Cu0.01% is below the Ag0.001%.Froth product-sulphur concentrate (sulphur, the sulfide and the gangue of copper, zinc etc.) is with soaking ore deposit agent molten soaking under 90 ℃ of left and right sides high temperature, and further behind recovery lead, silver, copper and the zinc wherein, final sulphur concentrate contains Pb＜1%.Ore dressing water is reused, and because of the ore pulp subacidity, floatation equipment must be anticorrosion.
Three, the electrolysis of the lead chloride after the flotation:, make catholyte with the sodium chloride aqueous solution dissolving of 4mol obtain purified lead chloride through flotation; Conversion fluid after reclaiming valuable metal is an anolyte.Carry out in the electrolyzer that the cathode ion exchange film is installed, the titanium sheet is made negative electrode, is coated with ruthenium (or ruthenium, iridium) titanium net and makes anode.
Molten again flotation lead chloride then recycles poor lead electrode liquid with electrolyzer behind the electrolysis one-period, under 50 ℃ of following temperature and mechanical stirring condition, be dissolved to more than the leaded 15g/l of solution, dense back overflowing liquid send electrolysis catholyte chamber's electrodeposition spongy lead, after thickened underflow further dissolves and washes away, make lead-in-dreg＜1% abandon it.Remove or partly remove ferrous chloride solution, adjustment PH to 1～1.5 of lead, silver, copper, zinc, be heated to 50～55 ℃, make anolyte, be sent to the electrolytic anode chamber and be oxidized to iron trichloride, make and soak ore deposit agent regeneration.
The anionic membrane press strip is embedded in makes barrier film on the plastic frame, be installed in the electrolyzer, and electrolytic anolyte compartment is separated with cathode compartment, and anolyte compartment's liquid level should be lower than the cathode compartment liquid level slightly.Because the effect of anionic membrane, electrolytic solution does not need deep purifying, just can the output lead bullion.
Electrolyzer is made by rigid plastic sheet, and an end makes sloping type.Bottom land is installed conveying belt, the spongy lead that the negative electrode electrodeposition goes out, is shed on the conveying belt, outside the output magazine.After the spongy lead washing densification, founding becomes lead pig.
The condition of electrolytic process control:
The composition of flotation lead chloride (%) is more than the Pb65, Cu0.01; Below the Ag0.001, catholyte composition (g/l) Cl150～170, Cu＜0.001, Ag＜0.001, Pb are not less than 9, anolyte composition (g/l) Pb＜5, Cu＜0.2, Ag＜0.001, Cl
-＞200, Fe
+ 2Be not less than 10.
Circulation of elecrolyte amount: 500～750ml/Ah, 45～50 ℃ of electrolyte temperatures, same pole span: 170～180mm, anode electrolysis density 250～600A/m
2, groove is pressed＜4.0 volts.The plumbous quality of electricity reaches more than the GB secondary.
Lead ore concentrate solid phase of the present invention transforms, the flow process of the wet method refining splicer skill of the selecting smelting combination of flotation removal of impurities, the diaphragm electrolysis lead dichloride aqueous solution, and shown in Figure of description: the S-among the figure represents solid phase, and L-represents water.
As can be seen from the figure, in whole flow process, water is boundary with the anion-exchange membrane, forms separately independently the anode recycle system and the negative electrode recycle system.Contain liquid before the approaching saturated negative electrode of lead chloride, enter the cathode compartment electrodeposition, the output spongy lead, a founding becomes lead pig through the group of pressure.Liquid is used to dissolve purified lead chloride behind the negative electrode, and insoluble gangue component is abandoned it.
The solid phase conversion fluid of concentrate of lead sulfide ore enters anolyte compartment's oxidation regeneration iron trichloride as liquid before the anode after reduction, iron filings or lead powder displacement, use for the solid phase conversion cycles.In this anolyte system, except that zinc is reclaimed in extraction, valuable metals such as silver, copper have also been reclaimed respectively.
Advantage of the present invention is:
(1), technical process is simple and easy, uses conventional hydrometallurgical plant, do not need special material anticorrosion, main process is operation under the condition of normal pressure and 60 ℃ all.Equipment there is not particular requirement.
(2), less energy consumption: solid phase transforms, reduction and displacement are all few in 60 ℃ of following operations, thermosteresis.Ore dressing also is to carry out at normal temperatures, transform slag flotation make meet the electrolysis requirement lead chloride with the sulphur concentrate, compare with additive method, process is simple, expense is few, energy consumption is low, does not have the gradient of temperature of crystallisation process.Because the operation solid-to-liquid ratio is little, the required power of circulate soln is few.
(3), good, this flow process of environmental protection is closed circuit.Do not use or produce any obnoxious flavour, three-waste free discharge, thereby can fundamentally solve the topsoil that exists in traditional pyrometallurgical smelting of lead and the hard nut to crack of poisoning.In addition, service temperature is not high, labor condition is good.
(4), the quality product height, the electricity that adopts this flow process can be directly to produce more than the GB secondary from lead ore concentrate is plumbous.Other foreign matter contents, can satisfy in the world the specification of quality of corrosion level metallic lead.
(5), the valuable metal of association in the raw material, silver, copper, zinc etc. can obtain enrichment largely in this flow process, and make pure metal or compound with general known method, thereby are comprehensively reclaimed.
(6), the main raw that cost is low, flow process is required is sodium-chlor and hydrochloric acid.Simple owing to flow process, as not have acid-base neutralisation process, materials consumption are mainly used in the loss that replenishes mechanical entrapment, and materials consumption seldom.
In a word, technical process of the present invention, treatment scale is changeable, and good adaptability is arranged.The wet method lead smelting process of this invention can be tightly according to mine, smelting on the spot, and the centering small mines is highly beneficial.In addition, also be suitable for handling based on lead, siliceous low Polymetallic sulphide concentrate.
Embodiment one
With Mt. Huang in Anhui Cen lead ore concentrate is raw material, lead ore concentrate chemical ingredients (weight ratio %):
Pb Ag Cu Zn S Fe Sio
2
62.25????0.075????0.94????4.50????15.77????4.70????5.01
Lead ore concentrate is finely ground to more than 90% by 0.055mm, and 10 kilograms of lead ore concentrates carry out solid phase and transform with soaking the ore deposit agent, and the composition that soaks the ore deposit agent is (g/l):
Fe
EntirelyFe
+ 3Cl
-Hcl
82.4????69.1????245????3.5
60 ℃ of controlled temperature, liquid-solid ratio=6: 1, mechanical stirring, transformation time 2 hours, liquid-solid separation back transforms slag (filter cake) with 5 premium on currency drip washing, obtain after the drip washing 11.73 kilograms of conversion slags, send to the flotation removal of impurities.The high ferro that transforms in the liquid of back is reduced into ferrous, 55 ℃ of reduction process controlled temperature, mechanical stirring, input concentrate amount and recovery time to control till high ferro fully is reduced in the liquid with lead ore concentrate to be transformed, lead sulfide is converted into solid-phase chlorination lead in reduction and conversion process, copper, zinc, silver etc. change solution over to, lead chloride reaches capacity in the solution simultaneously, and liquid-solid separation back reducing slag returns conversion.Reduced liquid at first is cooled to 15 ℃, and crystallization goes out lead chloride, uses silver in the iron powder replacement solution, copper and TBP extraction to reclaim zinc then, has got rid of ferrous chloride solution plumbous, silver-colored, copper zinc and has sent electrolysis to make anolyte.
Conversion slag after the drip washing carries out the flotation removal of impurities in flotation machine, pulp density 15%, and PH=2.5 through one roughing and once purging selection, obtains 7.99 kilograms of flotation lead chlorides.Its composition (weight ratio %):
Pb Ag Cu Sio
2
67.37????0.0005????0.007????4.90
Roughly select concentrate after primary cleaning, get 3.73 kilograms of flotation concentrates.Its composition (weight ratio %):
Pb????Ag????Cu
14.78????0.04????2.18
Further, finally discharge molten 2.37 kilograms of the slags that soak of system, composition (%) with after soaking that the ore deposit agent is molten and soaking:
Pb????Ag????Cu
0.90????0.029????2.28
Lead chloride after flotation is purified, evenly add in the mechanical agitating tank, the poor lead electrode liquid dissolving of the 4mol sodium-chlor that recycles with electrolyzer, 55～58 ℃ of controlled temperature, make that lead reaches more than the 13g/l in the solution, by adding the poly-propionic acid amide flocculation agent of 50,000 of lead chloride amount/adding, make insoluble white residue quicken sedimentation simultaneously.The dissolving tank overflow is dense to oblique plate closed box, and dense overflow clear liquid is sent to header tank, is circulated to electrolyzer and makes electrolysis cathode liquid.
Electrolysis is used S being equipped with
203Anion-exchange membrane makes in the membranous electrolyzer and carries out.Barrier film internal recycle anolyte, barrier film outer circulation catholyte.The titanium plate is made negative electrode, is coated with the ruthenium net and makes anode, with pole span 130mm, current density 300A/m
2Electrolyte temperature: 45～50 ℃, flow of electrolyte 800ml/Hhr.Anolyte compartment's ferrous chloride solution oxide becomes liquor ferri trichloridi.Make and soak ore deposit agent regeneration.Separate out spongy lead on the negative electrode.Electrolyzer one end makes sloping type, bottom land is installed and washed the conveying belt that Lun Bu makes, and the negative electrode spongy lead gathers certain thickness and comes off.Delivered to outside the groove by conveying belt, topped down at caustic soda after the spongy lead washing densification, founding becomes lead pig.Faradaic current efficient 94.9%, average groove is pressed 3.46 volts.Electricity lead composition (weight ratio %):
Ag????Cu????Zn????Fe????Bi
0.00085????0.0056????0.0001????0.0001????0.0011
As????Sn????Sb
0.0001????0.0001????0.0001
After the board concentrating box underflow dissolves and washes away with poor lead electrode liquid, 0.45 kilogram of final insoluble silicon.Slag ingredient (weight ratio %):
Pb Ag Cu Fe Sio
2
0.87????0.0004????0.019????7.59????40.96
Embodiment two:
With the Xi Tieshan of Qinghai Province lead ore concentrate is raw material, lead ore concentrate composition (weight ratio %):
Pb????Ag????Cu????Zn????S????Fe
64.80????0.058????0.23????2.5????11.19????6.07
Sio
2
0.39
10 kilograms of lead ore concentrates are finely ground to 92% and cross 0.0397mm, adopt the condition of embodiment one, will soak that hydrochloric acid increases to 8 grams per liters (because of the Alkaline minerals existence is arranged) in the agent of ore deposit, obtain 7.75 kilograms of flotation lead chlorides.Flotation lead chloride composition (weight ratio %)
Pb????Ag????Cu????Zn????Fe
72.39????0.0005????0.0052????0.026????0.09
Flotation concentrate is molten soak after, molten 2.866 kilograms of the quantity of slag of soaking.The molten composition that soaks slag:
Pb????Ag????Cu????S
0.83????0.029????2.28????41.8
The electrolysis of flotation lead chloride, current efficiency 95.63, average cell voltage=3.61 volt, electric lead composition weight percent (%):
Ag????Cu????Zn????Fe????Bi
0.00083????0.0075????0.0001????0.0001????0.0005
As????Sn????Sb
0.0001????0.0001????0.0001
0.14 kilogram of the insoluble white residue of output, weight ratio of constituents (%):
Pb Ag Cu Fe Sio
2
3.67????0.0013????0.042????8.06????45.6
Embodiment three:
Adopt the sub-lead ore in Qincheng city, Liaoning, its composition weight ratio (%):
Pb????Ag????Cu????Zn????Fe
67.64????0.1238????0.13????1.05????4.16
0.4 kilogram lead ore concentrate solid phase transforms, ore deposit agent composition (g/l) is soaked in control:
Fe
EntirelyFe
+ 2Cl Hcl
100????20????240????4
Liquid-solidization=5: 1,60 ℃ of temperature, transformation time 2 hours, after the liquid-solid separation, transform slag and in your sodium-chlor liquid of 4 mills, add 0.6 gram lead powder, adjust the flotation performance of silver, process control liquid-solid ratio=4.5,60 ℃ of temperature, PH=1, mechanical stirring 2 hours, liquid-solid separation back is adjusted slag and is sent the flotation removal of impurities, control pulp density 25%, PH=3, after roughly selecting, add 2 2# oil (terpenol) and 0.02 gram xanthate in the groove, carry out once purging selection and obtain flotation lead chloride 0.332kg.Flotation
Lead chloride weight ratio of constituents (%):
Pb????Ag????Cu
13.13????0.0006????0.0023
The flotation lead chloride dissolves with catholyte, molten 0.11 kilogram of the slag that soaks of output, and composition is weight ratio (%):
Pb????Ag????Cu
0.87????0.011????0.1
Control current density 500A/m
2, groove is pressed 3.7 volts, current efficiency 94.72%, and electric plumbous composition weight ratio (%):
Ag????Cu????Bi????Fe????As
0.00099????0.00054????0.00052????0.0002????0.0003
Zn????Sb????Sn
0.0003????0.0003????0.0003
0.005 kilogram of the insoluble silicon of output, its weight ratio of constituents (%):
Pb????Ag????Cu
0.91????0.01????0.008
Claims (3)
1, a kind of hydrometallurgical processes that from leaded sulphide ores, reclaims metallic lead, it is characterized in that vulcanizing lead ore and soaking ore deposit agent (iron trichloride, sodium chloride solution) carrying out solid phase in transforms, transform slag (lead chloride thereby make, elementary sulfur etc.) and conversion fluid (ferrous chloride etc.), transform slag through the flotation removal of impurities, obtain purified lead chloride, the aqueous solution with sodium-chlor is made catholyte to the lead chloride dissolving, conversion fluid is after reclaiming valuable metal wherein, be used as anolyte, adopt barrier film (anion-exchange membrane) electrolysis, deposit spongy lead from negative electrode, poor lead electrode liquid returns dissolving tank, have the recycle system of one's own, Fe on the anode
+ 2Oxidation generates Fe
+ 3Move to anolyte compartment and Fe with the chlorion that dissociates from cathode compartment through anion-exchange membrane
+ 3Generate iron trichloride, make and soak ore deposit agent regeneration,
The technical qualification of technological process:
(1) solid phase transforms:
(1) soak the composition of ore deposit agent:
Fe
3+： 40～80g/l PH＜1
Cl
-： 200～250g/l
Liquid-solid ratio: 4～8: 1; Temperature: 25～75 ℃
Stir: 1～4 hour; Transformation time: 1～4 hour
(2) transform the slag flotation removal of impurities:
Ore pulp (conversion slag) concentration (weight ratio): 10～25% PH=2～3
(3) diaphragm electrolysis:
Catholyte composition (g/l)
Cl
-：150～170； Cu＜0.001
Ag＜0.001；??????Pb＞9
Circulation of elecrolyte amount: 500～750ml/Ah
Electrolyte temperature: 45～50 ℃
Anodic current density: 250～600A/m
2
2, said according to claim 1, from leaded vulcanised ore concentrate, reclaim the hydrometallurgical processes of metallic lead, it is characterized in that adjusting the form that solid phase transforms silver in the slag with lead powder or other reductive agents, mine tailing (lead chloride) argentiferous that makes the flotation gained is below 10 gram/tons.
3, said according to claim 1 or 2, from leaded vulcanised ore concentrate, reclaim the hydrometallurgical processes of metallic lead, it is characterized in that catholyte is with the sodium chloride aqueous solution dissolving lead chloride of 4mol and get.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN88105063A CN1014722B (en) | 1988-05-17 | 1988-05-17 | New technology of wet lead-smelting |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN88105063A CN1014722B (en) | 1988-05-17 | 1988-05-17 | New technology of wet lead-smelting |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| CN1037739A CN1037739A (en) | 1989-12-06 |
| CN1014722B true CN1014722B (en) | 1991-11-13 |
Family
ID=4833294
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CN88105063A Expired CN1014722B (en) | 1988-05-17 | 1988-05-17 | New technology of wet lead-smelting |
Country Status (1)
| Country | Link |
|---|---|
| CN (1) | CN1014722B (en) |
Families Citing this family (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101619389B (en) * | 2009-08-03 | 2013-12-04 | 沈阳有色金属研究院 | Method for extracting lead from lead sulfide ore concentrate by HCl-O2-NaCl system |
-
1988
- 1988-05-17 CN CN88105063A patent/CN1014722B/en not_active Expired
Also Published As
| Publication number | Publication date |
|---|---|
| CN1037739A (en) | 1989-12-06 |
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