CN105964408A - Flotation method for recovering lead sulphate from lead-silver slag - Google Patents
Flotation method for recovering lead sulphate from lead-silver slag Download PDFInfo
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- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
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- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03B—SEPARATING SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS
- B03B1/00—Conditioning for facilitating separation by altering physical properties of the matter to be treated
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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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- 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
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- 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 invention relates to a flotation method for recovering lead sulphate from lead-silver slag. The method comprises the steps of (1) pulping condensation pretreatment, namely pulping the lead-silver slag with clear water, and condensing the pulp to a ground floor pulp concentration of 27-35%; (2) first coarse roughing, namely adding a pH adjusting agent, an inhibitor, a collector and a foaming agent to a floating ore-feeding in order for coarse roughing; (3) second sweeping selection, namely adding an inhibitor and a collector to the coarse roughing mine tailing in order for sweeping selection 1, and adding the inhibitor and the collector to the mine tailing obtained from sweeping selection 1 in order for sweeping selection 2, and the mine tailing obtained from the sweeping selection 2 is the final mine tailing; (4) third careful selection, namely adding the inhibitor to the coarse roughing concentrate for careful selection 1, and adding the inhibitor to the concentrate obtained from careful selection 1 for careful selection 2, the concentrate obtained from careful selection 1 enters careful selection 3 directly, and the concentrate obtained from careful selection 3 is the final concentrate. The method has the advantages of simple process flow, little equipment investment, high operationality, high lead grade and recovery rate in the concentrate, low production cost, great economic benefit, and obvious environment benefit.
Description
Technical field
The invention belongs to mineral production and processing and comprehensive reutilization technical field, be specifically related to one and return from lead smelting gas
Receive the method for floating of lead sulfate.
Background technology
Filtration mine tailing after zinc concentrate in oxygen pressure acid leaching slag flotation valuable element sulfur rich in lead, silver, therefore filter mine tailing also known as
For lead smelting gas.Owing to Leaching Systems is complex, leached mud composition is the most sufficiently complex, and the physico-chemical property of these compositions and nature
The mineral on boundary have had very big difference.At present, reclaim the main method of valuable metal composition in lead smelting gas and include solvent extraction method, pyrogenic process smelting
Refining method, floatation etc..The most complicated with the technological process of valuable metal in solvent extraction method recovery lead smelting gas, extraction time is relatively
Long, and leaching process needs to add the medicaments such as a large amount of strong acid, highly basic, higher to equipment requirements.Pyrometallurgical smelting method energy consumption is higher, and
New leaded slag, flue gas and flue dust can be produced, to environmentally undesirable.Flotation rule primarily focuses on the recovery of silver in lead smelting gas,
And the response rate of silver is relatively low, the recovery to lead is then studied less.
Currently, solvent extraction method reclaims the lead in lead smelting gas, main employing chlorination leaching lead method and alkali leaching leaching two kinds of leaching sides of lead method
Method.Chlorination leaching lead method products obtained therefrom is lead chloride, and by crystal form output, percent crystallization in massecuite is low, and product quality grade is low and purposes
Narrower；Alkali leaching leaching lead method typically uses the heavy lead of sulfuration, and vulcanized lead is only as metallurgical raw material.Pyrometallurgical smelting method can reclaim lead silver
Major part valuable metal in slag, but energy consumption is high, and environmental pollution is big etc., and problem fails effectively to solve.Therefore, a kind of work how is found
Process flow is simple, less to environmental hazard, with low cost, and what resource utilization was high reclaims the side of valuable metal composition in lead smelting gas
Method is problem demanding prompt solution.
Liu Canglong etc. (Liu Canglong, He Leqiong, Yang Shouming etc. process for recycling lead and silver by lead silver slag pyrogenic process.
CN101497944A), after lead smelting gas, industrial sodium carbonate, anthracite, iron filings being mixed by a certain percentage, it is heated up to 400-500 DEG C,
Keep 2.5-5h, continue heating after stirring, continue to be heated up to 850-950 DEG C, stop heating, carry out slag hitting, be cooled to
400-450 DEG C, carry out ingot casting and get final product.This technological temperature is high, and warm-up time is long, and therefore energy consumption is high, inevitable during heating
Meeting produce new flue gas, flue dust, slag etc..
Li Liting (Li Liting. utilize lead smelting gas comprehensively to extract the experimental study [J] of Zn, Pb and Ag. mineral products comprehensively utilize,
2010,3 (6): 15-18.) have studied the wet processing of employing " washing dezincification oxidizing roasting cyaniding leaching silver chlorination leaching lead "
The lead smelting gas leaching zinc hydrometallurgy carries out comprehensive reutilization.Silver leaching with cyanide adds environmentally harmful cyanide,
Chlorination leaching lead product is then lead chloride, and products application face is narrower.
Zhang Chuanbao, Yan Wenbin, Xu Hui etc. (Zhang Chuanbao, Yan Wenbin, Xu Hui etc. difficult Pb-Zn deposits acid leaching residue reclaims sulphuric acid
The technical study [J] of lead. application chemical industry, 2012,41 (7): 1188-1192.) use sodium chloride-sulfuric acid mixed solution to acidleach
Slag leaches, and is 33g/L at sodium chloride concentration, and liquid-solid ratio is 7:1, and the time is 1.5 h, temperature 95 DEG C, and sulfuric acid concentration is 1
Under conditions of mol/L, the leaching rate of lead is 82. 1%.Leaching process adds sulphuric acid and heats, higher for equipment requirements,
During heating, energy consumption is big, it addition, the leaching rate of lead is low.
Zhang Shaobo, Chen Wen, Shen Qianghua (Zhang Shaobo, Chen Wen, Shen Qianghua. from zinc leaching residue, obtain the test of lead silver preparation concentrate
Research [J]. mining and metallurgy, 2014,23 (1): 34-36.) for certain zinc leaching residue, after hot acid leaches, by one roughing, twice
Selected, that three times are scanned open circuit flow process flotation, in concentrate, silver grade can reach 2017.45g/t, and the response rate reaches
About 43.70%, in concentrate, lead grade reaches 50.81%.But in floatation process, add sodium sulfide, for bad environmental.
Li Zhengming, Zhang Wei, Dou Chuanlong etc. (Li Zhengming, Zhang Wei, Dou Chuanlong etc. in zinc hydrometallurgy, the process of lead smelting gas is reclaimed
Technique [J]. Yunnan metallurgy, 2011,40 (S): 173 175.) comprehensive method for floating of discussing processes lead smelting gas, it is indicated that domestic passes through
Lead smelting gas is reclaimed by flotation, and emphasis is the recovery of silver, and the response rate is low, and the research to lead is the most less.
In sum, in existing recovery lead silver there is energy consumption height in valuable metal method, and environmental pollution is big, to equipment requirements
The shortcomings such as height, leaching rate or the response rate are low.
Summary of the invention
It is an object of the invention to the deficiency existed for prior art, it is provided that a kind of from lead smelting gas, reclaim the floating of lead sulfate
Choosing method.
For achieving the above object, the technical solution used in the present invention is: a kind of flotation reclaiming lead sulfate from lead smelting gas
Method, comprise the steps be:
(1) slurrying concentrates pretreatment: with clear water by lead smelting gas slurrying, liquid-solid mass ratio is 8:1-12:1, concentrates ore pulp to bottom
Pulp density is 55-50%；Bottom ore pulp adds clear water slurrying, and liquid-solid mass ratio is 15:1-20:1, concentrates ore pulp to bottom
Pulp density is 27-35%；Obtained by bottom ore pulp as flotation to ore deposit；
(2) one roughing: pH regulator, inhibitor, collecting are added in the flotation obtained by step (1) the most successively to ore deposit
Agent and foaming agent, then roughly select, and roughly selects number of times for once；Rougher tailings enters (3rd) step, and rougher concentration enters (4th)
Step；
(3) scan for twice: the rougher tailings in step (2) adds inhibitor the most successively and collecting agent carries out scanning one,
Scan a gained concentrate and be back to roughly select to ore deposit, add inhibitor the most successively and collecting agent is carried out to scanning a gained mine tailing
Scan two, scan two gained concentrate return scan one give ore deposit, scanning two gained mine tailings is true tailings；
(4) triple cleaning: the rougher concentration in step (2) adds inhibitor and carries out selected, and a selected gained mine tailing returns
To roughly selecting to ore deposit, add inhibitor to a selected gained concentrate and enter selected two, selected two gained mine tailings be back to selected one to
Ore deposit, selected two gained concentrate are directly entered selected three, and selected three gained mine tailings are back to selected two to ore deposit, selected three gained concentrate
For final concentrate.
Described lead smelting gas is the filtration mine tailing after certain smeltery's oxygen pressure acid leaching residue flotation elementary sulfur, and filtering mine tailing is oxygen pressure acid
Waste residue after leaching zinc concentrate flotation elementary sulfur, lead grade is 17.5-19%.
Described pH regulator is calcium oxide, or is sodium hydroxide, or is sodium carbonate；Described inhibitor is waterglass；Described
Collecting agent is Salicyl Hydroximic Acid；Described foaming agent is terpenic oil.
It is 300-800g/t that described step (2) roughly selects middle pH regulator consumption, and regulation pH values of pulp is to 5.5-6.5, inhibitor
Consumption is 1500-2000g/t, and collector dosage is 1500-2000g/t, and frother dosages is 50-75g/t, and aeration quantity is
160-200L/h, flotation time is no less than 4min.
It is 500-1000g/t that described step (3) scans inhibitor consumption in, and collector dosage is 500-750g/t, fills
Tolerance is 160-200L/h, and flotation time is no less than 4min；Described to scan inhibitor consumption in two be 200-300g/t, collecting
Agent consumption is 150-250g/t, and aeration quantity is 160-200L/h, and flotation time is no less than 4min.
In described step (4) selected, inhibitor consumption is 500-750g/t, and aeration quantity is 160-200L/h, during flotation
Between no less than 4min；In described selected two, inhibitor consumption is 150-250g/t, and aeration quantity is 160-200L/h, flotation time
No less than 4min；Without any medicament in described selected three, aeration quantity is 160-200L/h, and flotation time is no less than 4min.
Filtration mine tailing clear water after described zinc concentrate in oxygen pressure acid leaching slag flotation elementary sulfur is by slurrying, and liquid-solid mass ratio is
10:1, ore pulp intermediate ion composition analysis result is: Zn2+, 1150 mg/l；Fe3+, 386 mg/l；Ca2+, 566 mg/l；Mg2+,
35.83 mg/l； Pb2+, 5.14 mg/l； Na+, 32.22 mg/l；K+, 7.46 mg/l； Cl-, 1.00 mg/l；SiO4 2-,
2.15 mg/l； SO4 2-, 4.5 g/l；Fe2+, 0.16 g/l；PH value is 2.25.
Owing to using such scheme, the present invention has the advantage that as follows:
1, the present invention uses slurrying to concentrate pre-pretreatment, decreases in the filtration mine tailing after oxygen pressure acid leaching residue flotation elementary sulfur unavoidable
Ion concentration, pH values of pulp rises；
2, the lead sulfate filtered in mine tailing after the present invention uses method for floating to reclaim oxygen pressure acid leaching residue flotation elementary sulfur, through once
Roughly selecting triple cleaning secondary to scan, in concentrate, lead grade is more than 50, and the response rate is more than 82；
3, the present invention uses method for floating to reclaim the lead sulfate in lead smelting gas, is effectively utilized by waste residue, remarkable in economical benefits, environment
Benefit is obvious；
4, the present invention uses method for floating to reclaim the lead sulfate in lead smelting gas, it is only necessary to interpolation flotation device, agitator, concentrator, if
Standby small investment, Factory Building takes up an area few；It addition, floating operation is convenient, simple, workable.
It is simple that the inventive method has technological process, and equipment investment is few, and Factory Building takes up an area few, workable, lead in concentrate
The advantages such as grade and the response rate are high, and production cost is low, remarkable in economical benefits, and environmental benefit is obvious.
Accompanying drawing explanation
Fig. 1 is the process chart of a kind of method for floating reclaiming lead sulfate from lead smelting gas.
Detailed description of the invention
The present invention is described further below in conjunction with the accompanying drawings.
The pH regulator calcium oxide or sodium hydroxide or the sodium carbonate that use in the embodiment of the present invention be, inhibitor water glass
Glass, collecting agent Salicyl Hydroximic Acid is analytical pure, and foaming agent terpenic oil is technical pure.
The embodiment of the present invention is roughly selected the HLXFD type flotation machine that floatation equipment is 1.5L of employing, scans the flotation of employing
Equipment is the HLXFD type flotation machine of 1.5L, and the floatation equipment of selected employing is the HLXFD type flotation machine of 0.75L.
Filtration mine tailing after the zinc concentrate in oxygen pressure acid leaching slag flotation elementary sulfur selected in the embodiment of the present invention is smelted from certain
Factory, fineness accounts for 90-93% for-0.074mm, and main gangue mineral is Gypsum Fibrosum and quartz, and main component contains by weight percentage
Pb17.5-19%, wherein the lead of 85-95% exists with lead sulfate form, containing Ag200-300g/t, containing SiO215-20%, containing Gypsum Fibrosum
25-35%。
Embodiment 1
Filtration mine tailing material fineness after zinc concentrate in oxygen pressure acid leaching slag flotation elementary sulfur accounts for 91.90% for-0.074mm, main component
By weight percentage containing Pb18.5%, containing Ag267g/t, containing SiO225%, containing Gypsum Fibrosum 30%.
The step of the method for floating reclaiming lead sulfate from raw material is:
(1) slurrying concentrates pretreatment: lead smelting gas slurrying, ore pulp agitator are stirred with the clear water that liquid-solid mass ratio is 10:1
5min, stands clarification, and the ore pulp siphon siphon supernatant clarified is 50% to bottom pulp density；Bottom ore pulp adds again
Entering clear water slurrying, liquid-solid mass ratio is 20:1, stirs 5min, stands clarification, the ore pulp siphon supernatant clarified, concentrates ore pulp
It is 31.24% to bottom pulp density；Obtained by bottom ore pulp as flotation to ore deposit；
(2) one roughing: calcium oxide 500g/t is added in the flotation obtained by step (1) the most successively to ore deposit, regulates ore deposit
Slurry pH to 6.1；Waterglass 2000g/t；Salicyl Hydroximic Acid 1500 g/t；Terpenic oil 75g/t；Control aeration quantity is 160L/h；So
After roughly select, roughly select number of times for once, flotation time is 4min；Rougher tailings enters (3rd) step, and rougher concentration enters the
(4) step；
(3) scan for twice: the rougher tailings in step (2) adds waterglass 750g/t and Salicyl Hydroximic Acid the most successively
500g/t carries out scanning one, and the aeration quantity scanning is 160 L/h, and flotation time is 4min；Scan a gained concentrate to be back to
Roughly select to ore deposit, add waterglass 250g/t and Salicyl Hydroximic Acid 200g/t the most successively scan to scanning a gained mine tailing
Two, the aeration quantity scanning two is 160 L/h, and flotation time is 4min, scan two gained concentrate return scan one give ore deposit, scan two
Gained mine tailing is true tailings；
(4) triple cleaning: the rougher concentration in step (2) adds waterglass 750g/t and carries out selected, a selected aeration quantity
Being 160 L/h, flotation time is 4min；A selected gained mine tailing is back to roughly select to ore deposit, adds water to a selected gained concentrate
Glass 200g/t enters selected two, and selected two aeration quantitys are 160 L/h, and flotation time is 4min；Selected two gained mine tailings return
Giving ore deposit to selected one, selected two gained concentrate are directly entered selected three without any medicament, and selected three aeration quantitys are 160L/h,
Flotation time is 4min；Selected three gained mine tailings be back to selected two give ore deposit, selected three gained concentrate are final concentrate.
In the final concentrate obtained, lead grade is 51.04%, and final concentrate lead recovery is 82.20%.
Embodiment 2
Filtration mine tailing raw material after the zinc concentrate in oxygen pressure acid leaching slag flotation elementary sulfur that the present embodiment is used is with embodiment 1.
The step of the method for floating reclaiming lead sulfate from raw material is:
(1) slurrying concentrates pretreatment: lead smelting gas slurrying, ore pulp agitator are stirred with the clear water that liquid-solid mass ratio is 12:1
5min, stands clarification, and the ore pulp siphon siphon supernatant clarified is 52% to bottom pulp density；Bottom ore pulp adds again
Entering clear water slurrying, liquid-solid mass ratio is 18:1, stirs 5min, stands clarification, the ore pulp siphon supernatant clarified, concentrates ore pulp
It is 30.23% to bottom pulp density；Obtained by bottom ore pulp as flotation to ore deposit；
(2) one roughing: sodium hydroxide 500g/t is added in the flotation obtained by step (1) the most successively to ore deposit, regulation
PH values of pulp is to 6.5；Waterglass 2000g/t；Salicyl Hydroximic Acid 1500g/t；Terpenic oil 75g/t；Control aeration quantity is 160L/h；So
After roughly select, roughly select number of times for once, flotation time is 4min.；Rougher tailings enters (3rd) step, and rougher concentration enters the
(4) step；
(3) scan for twice: the rougher tailings in step (2) adds waterglass 750g/t and Salicyl Hydroximic Acid the most successively
500g/t carries out scanning one, and the aeration quantity scanning is 160 L/h, and flotation time is 4min；Scan a gained concentrate to be back to
Roughly select to ore deposit, add waterglass 200g/t and Salicyl Hydroximic Acid 250g/t the most successively scan to scanning a gained mine tailing
Two, the aeration quantity scanning two is 160 L/h, and flotation time is 4min, scan two gained concentrate return scan one give ore deposit, scan two
Gained mine tailing is true tailings；
(4) triple cleaning: the rougher concentration in step (2) adds waterglass 750g/t and carries out selected, a selected aeration quantity
Being 160 L/h, flotation time is 4min；A selected gained mine tailing is back to roughly select to ore deposit, adds water to a selected gained concentrate
Glass 250g/t enters selected two, and selected two aeration quantitys are 160 L/h, and flotation time is 4min；Selected two gained mine tailings return
Giving ore deposit to selected one, selected two gained concentrate are directly entered selected three without any medicament, and selected three aeration quantitys are 160L/h,
Flotation time is 4min；Selected three gained mine tailings be back to selected two give ore deposit, selected three gained concentrate are final concentrate.
In the final concentrate obtained, lead grade is 50.43%, and final concentrate lead recovery is 81.03%.
Embodiment 3
Filtration mine tailing raw material after the zinc concentrate in oxygen pressure acid leaching slag flotation elementary sulfur that the present embodiment is used is with embodiment 1.
The characterization step of the method for floating reclaiming lead sulfate from raw material is:
(1) slurrying concentrates pretreatment: lead smelting gas slurrying, ore pulp agitator are stirred with the clear water that liquid-solid mass ratio is 8:1
5min, stands clarification, and the ore pulp siphon siphon supernatant clarified is 55% to bottom pulp density；Bottom ore pulp adds again
Entering clear water slurrying, liquid-solid mass ratio is 20:1, stirs 5min, stands clarification, the ore pulp siphon supernatant clarified, concentrates ore pulp
It is 29.56% to bottom pulp density；Obtained by bottom ore pulp as flotation to ore deposit；
(2) one roughing: sodium carbonate 500g/t is added in the flotation obtained by step (1) the most successively to ore deposit, regulates ore deposit
Slurry pH to 5.9；Waterglass 2000g/t；Salicyl Hydroximic Acid 1500 g/t；Terpenic oil 75g/t；Controlling aeration quantity is 160 L/h；So
After roughly select, roughly select number of times for once, flotation time is 4min.；Rougher tailings enters (3rd) step, and rougher concentration enters the
(4) step；
(3) scan for twice: the rougher tailings in step (2) adds waterglass 500g/t and Salicyl Hydroximic Acid the most successively
750g/t carries out scanning one, and the aeration quantity scanning is 160 L/h, and flotation time is 4min；Scan a gained concentrate to be back to
Roughly select to ore deposit, add waterglass 200g/t and Salicyl Hydroximic Acid 250g/t the most successively scan to scanning a gained mine tailing
Two, the aeration quantity scanning two is 160 L/h, and flotation time is 4min, scan two gained concentrate return scan one give ore deposit, scan two
Gained mine tailing is true tailings；
(4) triple cleaning: the rougher concentration in step (2) adds waterglass 750g/t and carries out selected, a selected aeration quantity
Being 160 L/h, flotation time is 4min；A selected gained mine tailing is back to roughly select to ore deposit, adds water to a selected gained concentrate
Glass 250g/t enters selected two, and selected two aeration quantitys are 160 L/h, and flotation time is 4min；Selected two gained mine tailings return
Giving ore deposit to selected one, selected two gained concentrate are directly entered selected three without any medicament, and selected three aeration quantitys are 160 L/
H, flotation time is 4min；Selected three gained mine tailings be back to selected two give ore deposit, selected three gained concentrate are final concentrate.
In the final concentrate obtained, lead grade is 50.28%, and final concentrate lead recovery is 81.11%.
Claims (5)
1. reclaiming a method for floating for lead sulfate from lead smelting gas, described lead smelting gas is that certain smeltery's oxygen pressure acid leaching residue flotation is first
Filtration mine tailing after element sulfur, wherein lead grade 17.5-19%, its characterization step is:
(1) slurrying concentrates pretreatment: with clear water by lead smelting gas slurrying, liquid-solid mass ratio is 8:1-12:1, concentrates ore pulp to bottom
Pulp density is 55-50%；Bottom ore pulp adds clear water slurrying, and liquid-solid mass ratio is 15:1-20:1, concentrates ore pulp to bottom
Pulp density is 27-35%；Obtained by bottom ore pulp as flotation to ore deposit；
(2) one roughing: pH regulator, inhibitor, collecting are added in the flotation obtained by step (1) the most successively to ore deposit
Agent and foaming agent, then roughly select, and roughly selects number of times for once；Rougher tailings enters (3rd) step, and rougher concentration enters (4th)
Step；
(3) scan for twice: the rougher tailings in step (2) adds inhibitor the most successively and collecting agent carries out scanning one,
Scan a gained concentrate and be back to roughly select to ore deposit, add inhibitor the most successively and collecting agent is carried out to scanning a gained mine tailing
Scan two, scan two gained concentrate return scan one give ore deposit, scanning two gained mine tailings is true tailings；
(4) triple cleaning: the rougher concentration in step (2) adds inhibitor and carries out selected, and a selected gained mine tailing returns
To roughly selecting to ore deposit, add inhibitor to a selected gained concentrate and enter selected two, selected two gained mine tailings be back to selected one to
Ore deposit, selected two gained concentrate are directly entered selected three, and selected three gained mine tailings are back to selected two to ore deposit, selected three gained concentrate
For final concentrate.
A kind of method for floating reclaiming lead sulfate from lead smelting gas, it is characterised in that: described pH adjusts
Whole dose is calcium oxide, or is sodium hydroxide, or is sodium carbonate；Described inhibitor is waterglass；Described collecting agent is bigcatkin willow hydroxyl oxime
Acid；Described foaming agent is terpenic oil.
A kind of method for floating reclaiming lead sulfate from lead smelting gas, it is characterised in that described step
(2) roughly selecting middle pH regulator consumption is 300-800g/t, regulation pH values of pulp to 5.5-6.5, inhibitor consumption is 1500-
2000g/t, collector dosage is 1500-2000g/t, and frother dosages is 50-75g/t, and aeration quantity is 160-200L/h,
Flotation time is no less than 4min.
A kind of method for floating reclaiming lead sulfate from lead smelting gas, it is characterised in that described step
(3) scanning inhibitor consumption in is 500-1000g/t, and collector dosage is 500-750g/t, and aeration quantity is 160-200L/
H, flotation time is no less than 4min；Described to scan inhibitor consumption in two be 200-300g/t, and collector dosage is 150-250g/
T, aeration quantity is 160-200L/h, and flotation time is no less than 4min.
A kind of method for floating reclaiming lead sulfate from lead smelting gas, it is characterised in that described step
(4) inhibitor consumption is 500-750g/t in selected one, and aeration quantity is 160-200L/h, and flotation time is no less than 4min；Described
In selected two, inhibitor consumption is 150-250g/t, and aeration quantity is 160-200L/h, and flotation time is no less than 4min；Described essence
Selecting without any medicament in three, aeration quantity is 160-200L/h, and flotation time is no less than 4min.
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Cited By (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN109734057A (en) * | 2019-01-02 | 2019-05-10 | 长沙有色冶金设计研究院有限公司 | A kind of method of zinc oxygen leaching slag Recovered sulphur and silver |
| CN113926593A (en) * | 2021-09-30 | 2022-01-14 | 深圳市中金岭南有色金属股份有限公司凡口铅锌矿 | Lead-sulfur mixed floating method |
Citations (6)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| WO2008001945A1 (en) * | 2006-06-30 | 2008-01-03 | Nikon Corporation | Microscope device |
| CN103433120A (en) * | 2013-08-28 | 2013-12-11 | 武汉科技大学 | Heavy-floating joint separation method for producing high-grade sulfur concentrate |
| CN104722410A (en) * | 2015-02-16 | 2015-06-24 | 巴彦淖尔紫金有色金属有限公司 | Method for extracting and separating valuable metal from hot-acid leaching residues produced by wet-process zinc smelting |
| CN104841564A (en) * | 2015-05-15 | 2015-08-19 | 西北矿冶研究院 | Lead-silver residue flotation combined collecting agent and application process thereof |
| CN104846208A (en) * | 2015-05-15 | 2015-08-19 | 西北矿冶研究院 | Method for comprehensively recovering gold and silver in lead-silver slag |
| CN104907182A (en) * | 2015-05-15 | 2015-09-16 | 西北矿冶研究院 | Lead-silver slag flotation composite inhibitor and application process thereof |
-
2015
- 2015-11-02 CN CN201510729945.1A patent/CN105964408A/en active Pending
Patent Citations (6)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| WO2008001945A1 (en) * | 2006-06-30 | 2008-01-03 | Nikon Corporation | Microscope device |
| CN103433120A (en) * | 2013-08-28 | 2013-12-11 | 武汉科技大学 | Heavy-floating joint separation method for producing high-grade sulfur concentrate |
| CN104722410A (en) * | 2015-02-16 | 2015-06-24 | 巴彦淖尔紫金有色金属有限公司 | Method for extracting and separating valuable metal from hot-acid leaching residues produced by wet-process zinc smelting |
| CN104841564A (en) * | 2015-05-15 | 2015-08-19 | 西北矿冶研究院 | Lead-silver residue flotation combined collecting agent and application process thereof |
| CN104846208A (en) * | 2015-05-15 | 2015-08-19 | 西北矿冶研究院 | Method for comprehensively recovering gold and silver in lead-silver slag |
| CN104907182A (en) * | 2015-05-15 | 2015-09-16 | 西北矿冶研究院 | Lead-silver slag flotation composite inhibitor and application process thereof |
Non-Patent Citations (8)
| Title |
|---|
| F.拉什申等: ""铅矾浮选：从锌浸出渣中回收铅的研究"", 《国外金属选矿》 * |
| 张少博等: ""从锌浸出渣中获得铅银精矿的试验研究"", 《矿冶》 * |
| 李正明等: ""湿法炼锌中铅银渣的处理回收工艺"", 《云南冶金》 * |
| 李黎婷: ""利用铅银渣综合提取锌铅银的试验研究"", 《矿产综合利用》 * |
| 王树立等: ""铅银渣中多种元素提取工艺路线的研究"", 《现代化工》 * |
| 胡岳华等: "《矿物浮选》", 31 March 2014, 中南大学出版社 * |
| 赵景云等: ""水杨羟肟酸浮选菱锌矿和硫酸铅试验"", 《有色金属》 * |
| 陈卫华: ""硫化钠在锌浸出渣银浮选中的研究"", 《有色金属》 * |
Cited By (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN109734057A (en) * | 2019-01-02 | 2019-05-10 | 长沙有色冶金设计研究院有限公司 | A kind of method of zinc oxygen leaching slag Recovered sulphur and silver |
| CN113926593A (en) * | 2021-09-30 | 2022-01-14 | 深圳市中金岭南有色金属股份有限公司凡口铅锌矿 | Lead-sulfur mixed floating method |
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