CN116812874B - Method for efficiently recycling sulfur and zinc and silver from zinc hydrometallurgy high-sulfur residues - Google Patents
Method for efficiently recycling sulfur and zinc and silver from zinc hydrometallurgy high-sulfur residues Download PDFInfo
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- CN116812874B CN116812874B CN202311100902.8A CN202311100902A CN116812874B CN 116812874 B CN116812874 B CN 116812874B CN 202311100902 A CN202311100902 A CN 202311100902A CN 116812874 B CN116812874 B CN 116812874B
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- 229910052717 sulfur Inorganic materials 0.000 title claims abstract description 87
- 239000011593 sulfur Substances 0.000 title claims abstract description 87
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 title claims abstract description 86
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 title claims abstract description 59
- 229910052725 zinc Inorganic materials 0.000 title claims abstract description 59
- 239000011701 zinc Substances 0.000 title claims abstract description 59
- 229910052709 silver Inorganic materials 0.000 title claims abstract description 43
- 239000004332 silver Substances 0.000 title claims abstract description 43
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 title claims abstract description 42
- 238000000034 method Methods 0.000 title claims abstract description 22
- 238000009854 hydrometallurgy Methods 0.000 title claims abstract description 11
- 238000004064 recycling Methods 0.000 title claims abstract description 6
- 239000002893 slag Substances 0.000 claims abstract description 57
- 239000007788 liquid Substances 0.000 claims abstract description 21
- 238000005188 flotation Methods 0.000 claims abstract description 20
- IWLXWEWGQZEKGZ-UHFFFAOYSA-N azane;zinc Chemical compound N.[Zn] IWLXWEWGQZEKGZ-UHFFFAOYSA-N 0.000 claims abstract description 18
- 238000006243 chemical reaction Methods 0.000 claims abstract description 18
- 239000002002 slurry Substances 0.000 claims abstract description 17
- 239000003112 inhibitor Substances 0.000 claims abstract description 16
- VHUUQVKOLVNVRT-UHFFFAOYSA-N Ammonium hydroxide Chemical compound [NH4+].[OH-] VHUUQVKOLVNVRT-UHFFFAOYSA-N 0.000 claims abstract description 14
- CURLTUGMZLYLDI-UHFFFAOYSA-N Carbon dioxide Chemical compound O=C=O CURLTUGMZLYLDI-UHFFFAOYSA-N 0.000 claims abstract description 14
- 235000011114 ammonium hydroxide Nutrition 0.000 claims abstract description 14
- 238000002386 leaching Methods 0.000 claims abstract description 14
- 239000012141 concentrate Substances 0.000 claims abstract description 11
- UOURRHZRLGCVDA-UHFFFAOYSA-D pentazinc;dicarbonate;hexahydroxide Chemical compound [OH-].[OH-].[OH-].[OH-].[OH-].[OH-].[Zn+2].[Zn+2].[Zn+2].[Zn+2].[Zn+2].[O-]C([O-])=O.[O-]C([O-])=O UOURRHZRLGCVDA-UHFFFAOYSA-D 0.000 claims abstract description 10
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims abstract description 9
- 239000000411 inducer Substances 0.000 claims abstract description 9
- 229910052760 oxygen Inorganic materials 0.000 claims abstract description 9
- 239000001301 oxygen Substances 0.000 claims abstract description 9
- NFMAZVUSKIJEIH-UHFFFAOYSA-N bis(sulfanylidene)iron Chemical compound S=[Fe]=S NFMAZVUSKIJEIH-UHFFFAOYSA-N 0.000 claims abstract description 7
- 239000001569 carbon dioxide Substances 0.000 claims abstract description 7
- 229910002092 carbon dioxide Inorganic materials 0.000 claims abstract description 7
- 229910000339 iron disulfide Inorganic materials 0.000 claims abstract description 7
- RUTXIHLAWFEWGM-UHFFFAOYSA-H iron(3+) sulfate Chemical compound [Fe+3].[Fe+3].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O RUTXIHLAWFEWGM-UHFFFAOYSA-H 0.000 claims abstract description 7
- 229910000360 iron(III) sulfate Inorganic materials 0.000 claims abstract description 7
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 24
- 238000001914 filtration Methods 0.000 claims description 21
- 238000003756 stirring Methods 0.000 claims description 18
- GEHJYWRUCIMESM-UHFFFAOYSA-L sodium sulfite Chemical compound [Na+].[Na+].[O-]S([O-])=O GEHJYWRUCIMESM-UHFFFAOYSA-L 0.000 claims description 14
- WSFSSNUMVMOOMR-UHFFFAOYSA-N Formaldehyde Chemical compound O=C WSFSSNUMVMOOMR-UHFFFAOYSA-N 0.000 claims description 11
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims description 8
- 235000010265 sodium sulphite Nutrition 0.000 claims description 7
- NWONKYPBYAMBJT-UHFFFAOYSA-L zinc sulfate Chemical compound [Zn+2].[O-]S([O-])(=O)=O NWONKYPBYAMBJT-UHFFFAOYSA-L 0.000 claims description 7
- 229910000368 zinc sulfate Inorganic materials 0.000 claims description 7
- 229960001763 zinc sulfate Drugs 0.000 claims description 7
- IKHGUXGNUITLKF-UHFFFAOYSA-N Acetaldehyde Chemical compound CC=O IKHGUXGNUITLKF-UHFFFAOYSA-N 0.000 claims description 6
- PLKATZNSTYDYJW-UHFFFAOYSA-N azane silver Chemical compound N.[Ag] PLKATZNSTYDYJW-UHFFFAOYSA-N 0.000 claims description 6
- 239000007791 liquid phase Substances 0.000 claims description 5
- 239000002244 precipitate Substances 0.000 claims description 5
- 239000013049 sediment Substances 0.000 claims description 5
- 239000002912 waste gas Substances 0.000 claims description 5
- UYJXRRSPUVSSMN-UHFFFAOYSA-P ammonium sulfide Chemical compound [NH4+].[NH4+].[S-2] UYJXRRSPUVSSMN-UHFFFAOYSA-P 0.000 claims description 4
- 239000003814 drug Substances 0.000 claims description 3
- 229910052979 sodium sulfide Inorganic materials 0.000 claims description 3
- GRVFOGOEDUUMBP-UHFFFAOYSA-N sodium sulfide (anhydrous) Chemical compound [Na+].[Na+].[S-2] GRVFOGOEDUUMBP-UHFFFAOYSA-N 0.000 claims description 3
- 238000011084 recovery Methods 0.000 abstract description 44
- 229910052751 metal Inorganic materials 0.000 abstract description 21
- 239000002184 metal Substances 0.000 abstract description 19
- 150000002739 metals Chemical class 0.000 abstract description 11
- 239000000047 product Substances 0.000 abstract description 7
- 238000000746 purification Methods 0.000 abstract description 3
- 238000010668 complexation reaction Methods 0.000 abstract description 2
- 239000000706 filtrate Substances 0.000 abstract description 2
- 230000000536 complexating effect Effects 0.000 abstract 1
- 239000011133 lead Substances 0.000 description 13
- 238000003723 Smelting Methods 0.000 description 4
- 230000000694 effects Effects 0.000 description 4
- 239000002699 waste material Substances 0.000 description 4
- BFNBIHQBYMNNAN-UHFFFAOYSA-N ammonium sulfate Chemical compound N.N.OS(O)(=O)=O BFNBIHQBYMNNAN-UHFFFAOYSA-N 0.000 description 3
- 229910052921 ammonium sulfate Inorganic materials 0.000 description 3
- 235000011130 ammonium sulphate Nutrition 0.000 description 3
- 239000003795 chemical substances by application Substances 0.000 description 3
- 229910000358 iron sulfate Inorganic materials 0.000 description 3
- BAUYGSIQEAFULO-UHFFFAOYSA-L iron(2+) sulfate (anhydrous) Chemical compound [Fe+2].[O-]S([O-])(=O)=O BAUYGSIQEAFULO-UHFFFAOYSA-L 0.000 description 3
- BSWGGJHLVUUXTL-UHFFFAOYSA-N silver zinc Chemical compound [Zn].[Ag] BSWGGJHLVUUXTL-UHFFFAOYSA-N 0.000 description 3
- QAOWNCQODCNURD-UHFFFAOYSA-L Sulfate Chemical compound [O-]S([O-])(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-L 0.000 description 2
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 description 2
- XOCUXOWLYLLJLV-UHFFFAOYSA-N [O].[S] Chemical compound [O].[S] XOCUXOWLYLLJLV-UHFFFAOYSA-N 0.000 description 2
- 238000005516 engineering process Methods 0.000 description 2
- 230000007613 environmental effect Effects 0.000 description 2
- 239000008098 formaldehyde solution Substances 0.000 description 2
- SXUIEYVKFUGEOZ-UHFFFAOYSA-N N.[Zn].[Ag] Chemical compound N.[Zn].[Ag] SXUIEYVKFUGEOZ-UHFFFAOYSA-N 0.000 description 1
- 239000002253 acid Substances 0.000 description 1
- 229910052785 arsenic Inorganic materials 0.000 description 1
- RQNWIZPPADIBDY-UHFFFAOYSA-N arsenic atom Chemical compound [As] RQNWIZPPADIBDY-UHFFFAOYSA-N 0.000 description 1
- 238000007664 blowing Methods 0.000 description 1
- 229910052793 cadmium Inorganic materials 0.000 description 1
- BDOSMKKIYDKNTQ-UHFFFAOYSA-N cadmium atom Chemical compound [Cd] BDOSMKKIYDKNTQ-UHFFFAOYSA-N 0.000 description 1
- 239000013064 chemical raw material Substances 0.000 description 1
- 239000003153 chemical reaction reagent Substances 0.000 description 1
- 239000011248 coating agent Substances 0.000 description 1
- 238000000576 coating method Methods 0.000 description 1
- 230000000052 comparative effect Effects 0.000 description 1
- 238000001816 cooling Methods 0.000 description 1
- 238000011161 development Methods 0.000 description 1
- 238000001704 evaporation Methods 0.000 description 1
- 230000008020 evaporation Effects 0.000 description 1
- 229910052732 germanium Inorganic materials 0.000 description 1
- GNPVGFCGXDBREM-UHFFFAOYSA-N germanium atom Chemical compound [Ge] GNPVGFCGXDBREM-UHFFFAOYSA-N 0.000 description 1
- 229910001385 heavy metal Inorganic materials 0.000 description 1
- 238000007654 immersion Methods 0.000 description 1
- 239000012535 impurity Substances 0.000 description 1
- XEEYBQQBJWHFJM-UHFFFAOYSA-N iron Substances [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 1
- 238000004519 manufacturing process Methods 0.000 description 1
- QSHDDOUJBYECFT-UHFFFAOYSA-N mercury Chemical compound [Hg] QSHDDOUJBYECFT-UHFFFAOYSA-N 0.000 description 1
- 229910052753 mercury Inorganic materials 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- -1 organic synthesis Substances 0.000 description 1
- 239000012071 phase Substances 0.000 description 1
- 239000002994 raw material Substances 0.000 description 1
- 230000035484 reaction time Effects 0.000 description 1
- 238000011160 research Methods 0.000 description 1
- 239000007787 solid Substances 0.000 description 1
- WGPCGCOKHWGKJJ-UHFFFAOYSA-N sulfanylidenezinc Chemical compound [Zn]=S WGPCGCOKHWGKJJ-UHFFFAOYSA-N 0.000 description 1
- 150000004763 sulfides Chemical class 0.000 description 1
- 238000003786 synthesis reaction Methods 0.000 description 1
- 238000012546 transfer Methods 0.000 description 1
Classifications
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01B—NON-METALLIC ELEMENTS; COMPOUNDS THEREOF; METALLOIDS OR COMPOUNDS THEREOF NOT COVERED BY SUBCLASS C01C
- C01B17/00—Sulfur; Compounds thereof
- C01B17/02—Preparation of sulfur; Purification
- C01B17/06—Preparation of sulfur; Purification from non-gaseous sulfides or materials containing such sulfides, e.g. ores
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01G—COMPOUNDS CONTAINING METALS NOT COVERED BY SUBCLASSES C01D OR C01F
- C01G9/00—Compounds of zinc
-
- 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
- C22B11/00—Obtaining noble metals
- C22B11/04—Obtaining noble metals by wet processes
- C22B11/042—Recovery of noble metals from waste materials
-
- 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
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/006—Wet processes
-
- 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
Landscapes
- Chemical & Material Sciences (AREA)
- Organic Chemistry (AREA)
- Engineering & Computer Science (AREA)
- Geology (AREA)
- Inorganic Chemistry (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Life Sciences & Earth Sciences (AREA)
- Manufacturing & Machinery (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Metallurgy (AREA)
- Environmental & Geological Engineering (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention discloses a method for efficiently recycling sulfur and zinc and silver from zinc hydrometallurgy high-sulfur residues, belonging to the field of resource recycling. The oxygen pressure leaching second-stage underflow liquid contains a large amount of high-sulfur slag, and a proper amount of ferric sulfate is added as an inducer to convert iron disulfide in the high-sulfur slag into elemental sulfur, so that the content of the elemental sulfur is improved; controlling reaction conditions and adding inhibitors in the flotation stage, improving the recovery rate of sulfur, enabling more metals such as lead and zinc to enter tailings, and providing a basis for further improving the purity of the sulfur; mixing tailing slag and hot filtered slag with filtrate to prepare slurry, adding ammonia water into the slurry, dissolving zinc and silver by a zinc ammonia complexation method, recovering metal silver by adopting a silver mirror reaction, and finally absorbing and capturing carbon dioxide by using complexing liquid to prepare a basic zinc carbonate product. The method has simple operation and good application prospect, and solves the problems that the sulfur recovery rate is low, the sulfur purification is affected by lead and zinc entering concentrate, the zinc and silver recovery rate is low and the recovery cost is high in the existing process.
Description
Technical Field
The invention relates to the technical field of resource recovery, in particular to a method for efficiently recovering sulfur and zinc and silver from high-sulfur slag of zinc hydrometallurgy.
Background
Sulfur is used as a basic chemical raw material and widely applied to industries such as coating, organic synthesis, acid production, medicine, food and the like, and sulfur products are mainly from naturally occurring and recovered sulfur at present. The recovered sulfur has the advantages of high purity, less impurities, stable quality and the like. The sulfur demand of China is large each year, the sulfur output of China is small, and the external dependence is high.
The wet zinc smelting process can produce a large amount of leaching slag, and because the sulfur content is high, the sulfur grade is about 40-55 percent, and the sulfur mainly exists in the forms of elemental sulfur, sulfide, sulfate and the like, and contains valuable metal elements such as silver, lead, zinc, germanium and the like, the high sulfur slag is comprehensively recovered, and the sulfur and valuable metal recovery technology in the high sulfur slag by the development method has important significance in relieving contradiction between sulfur supply and demand in China, maximally improving the resource utilization level and relieving the environmental pressure caused by the piling of leaching slag. The traditional treatment method has low efficiency, causes resource waste and can produce secondary pollution. Therefore, how to efficiently and environmentally recover sulfur and valuable metals (such as zinc and silver) in high-sulfur slag is an important subject in the metallurgical industry and environmental protection fields today. Patent 20141068240. X reports a method for recovering sulfur from zinc leached sulfur-containing slag, which uses a high-pressure reaction kettle to transform the sulfur slag to improve the recovery rate of sulfur, but has higher cost and inconvenient operation. The 201310422340.9 patent firstly rapidly heats and boosts oxygen immersion liquid in normal pressure oxygen-enriched direct zinc hydrometallurgy, and carries out flash evaporation, then uses hot filtration to separate element sulfur, and then uses water cooling to melt the element sulfur into granular solid element sulfur, so that the recovery rate of sulfur element is difficult to ensure. Patent 202210102498.7 pulverizes and sieves the zinc-sulfur slag again and carries out secondary flotation to improve the recovery rate of sulfur concentrate, but does not consider the operation cost, recovery of valuable metal elements, and the like.
At present, the research on the recovery of sulfur from high-sulfur slag is more, but the technology of low cost, high recovery rate and recovery of valuable metal resources is still a difficult problem to overcome. Therefore, the invention develops a method for efficiently recovering sulfur and zinc and silver from the high-sulfur slag of zinc hydrometallurgy aiming at the bottleneck problem of recovering sulfur and valuable metals from the high-sulfur slag at present.
Disclosure of Invention
Aiming at the problems, the invention provides a method for efficiently recovering sulfur and zinc and silver from zinc hydrometallurgy high-sulfur slag, which solves the problems that the recovery rate of sulfur is low, the recovery rate of zinc and silver is low and the recovery cost is high due to the fact that lead and zinc enter concentrate to affect sulfur purification.
In order to achieve the above purpose, the present invention adopts the following technical scheme:
step one: adding an inducer into the underflow liquid of the second stage of oxygen pressure leaching, stirring, and converting iron disulfide in the high-sulfur slag into elemental sulfur;
step two: in the flotation stage, sodium hydroxide is added to adjust the pH value of the underflow, air is blown, an inhibitor is added, and flotation is carried out for 8-30 minutes;
step three: the concentrate obtained by floatation is conveyed to a hot filtration section to recycle high-purity sulfur, and hot filter residues are returned to the slurry;
step four: adding ammonia water into the slurry, stirring at a certain temperature, reacting to obtain zinc ammonia complex and silver ammonia complex, and entering into a liquid phase;
step five: filtering, adding formaldehyde or acetaldehyde solution into the water bath until no precipitate is generated in the solution, and recovering silver and zinc ammonia complex solution;
step six: the zinc ammonia complex solution captures and absorbs the waste gas containing carbon dioxide, reacts under certain conditions until the solution does not generate sediment any more, and is filtered to obtain basic zinc carbonate.
Preferably, the two-stage underflow liquid in the first step contains high-sulfur slag, the pH value is 1-3, the temperature is 60-80 ℃, and the solid-liquid ratio is 1:4-1:10.
Preferably, the inducer in the first step is ferric sulfate, and the concentration of the ferric sulfate in the underflow is 0.01mol/L to 0.5mol/L.
Preferably, in the first step, the stirring speed is 100-500 rpm, and the stirring time is 30-120 minutes.
The technical effect achieved by the first step is as follows: the ferric sulfate converts iron disulfide in the high-sulfur slag into elemental sulfur, so that the content of the elemental sulfur is improved, and the recovery rate is improved by 2% -8%.
The reaction principle is as follows:
FeS 2 +Fe 2 (SO 4 ) 3 =3FeSO 4 +2S
preferably, in the second step, the pH value is 1.5-4, and the air flow rate is 1-5L/min.
Preferably, the inhibitor in the second step is a combination of sodium sulfide, zinc sulfate and sodium sulfite, or a combination of ammonium sulfide, sodium hydroxide, zinc sulfate and sodium sulfite.
Preferably, the addition mass of each medicament in the inhibitor is the same, and the addition mass is 100-300 g/ton of slag.
The technical effect achieved by adopting the second step is as follows: the recovery rate of sulfur is improved by controlling the reaction conditions and the added inhibitor, and is more than 95 percent; and more metals such as lead, zinc and the like enter the tailing slag.
Preferably, the concentration of the ammonia water in the slurry in the fourth step is 0.5-2 mol/L;
the temperature is 20-60 ℃, the stirring speed is 100-400 rpm, and the stirring time is 1-3 hours.
Preferably, the water bath temperature in the fifth step is 50-70 ℃.
The technical effect that adopts above-mentioned technical scheme to reach is: the recovery rate of silver is 60-90%.
Preferably, the reaction temperature in the step six is 20-40 ℃ and the pH value is 4-7.
The technical effect that adopts above-mentioned technical scheme to reach is: the recovery rate of zinc is 65-90%.
The wet zinc smelting process can produce a large amount of leaching slag, and the leaching slag often contains various harmful elements such as lead, mercury, arsenic, cadmium and the like, is listed into HW48 nonferrous metal smelting waste, has great harm to the environment and human health, and meanwhile, the slag contains high sulfur content and various valuable metals such as zinc, lead, silver and the like, so the leaching slag is also one of important raw materials for recycling sulfur, silver, zinc and lead.
In order to efficiently separate heavy metal phases from elemental sulfur in the sulfur-oxygen containing pressure leaching slag and directionally recycle zinc and silver valuable metals in the form of sulfate in the sulfur-oxygen containing pressure leaching slag, the invention is convenient for recovery in a smelting process and overcomes the technical problem of high recovery cost in the prior art, and the technical conception of the invention is as follows:
considering that a part of sulfur elements in the high-sulfur slag exist in the form of sulfides, if the part of sulfur elements are converted into sulfur, the recovery rate of the sulfur is improved; since lead and zinc can influence the subsequent purification process of sulfur, the addition of inhibitors in the flotation stage allows more lead and zinc to enter the tailings; the slag and the filtrate still contain valuable metals, the zinc and the silver are recovered by adopting a step recovery method, the zinc and the silver are dissolved into the solution through a complexation reaction by combining the physicochemical characteristics of the silver and the zinc, then the metal silver is recovered through a silver mirror reaction, and the metal zinc is recovered through a carbonation process; the invention solves the problems of low sulfur recovery rate, low zinc and silver recovery rate and high recovery cost caused by lead and zinc entering concentrate in the prior art.
Compared with the prior art, the invention has the advantages that:
(1) The sulfide in the high-sulfur slag is converted into elemental sulfur by adding the ferric sulfate inducer, so that the recovery rate of sulfur is improved, and the operation is simple.
(2) The flotation conditions are controlled, and the inhibitor is added to improve the recovery rate of sulfur, inhibit the transfer of zinc and lead to concentrate, and reduce the influence on the subsequent recovery of sulfur.
(3) Zinc and silver in the residual slag are leached out in the form of zinc ammonia and silver ammonia complex by adopting a zinc-silver ammonia complex principle, metal silver is recovered by a silver mirror method, and carbon dioxide is absorbed and trapped by a zinc ammonia complex solution to prepare a basic zinc carbonate product, so that the metal zinc is efficiently recovered, and the purposes of treating waste by waste, saving energy and reducing emission are realized.
Detailed Description
The following description of the embodiments of the present invention will be made clearly and completely, and it is apparent that the described embodiments are only some embodiments of the present invention, but not all embodiments. All other embodiments, which can be made by those skilled in the art based on the embodiments of the invention without making any inventive effort, are intended to be within the scope of the invention.
The following examples are not intended to be limiting, and the reagents used are all commercially available products or are prepared by conventional means using equipment conventional in the art, and the invention is further described in connection with the specific examples.
Example 1
Step one: taking 1L of oxygen pressure leaching second-stage underflow liquid, wherein the underflow liquid contains high sulfur slag, the pH value of the solution is 1, the temperature of the solution is 80 ℃, and the solid-liquid ratio (volume ratio) is 1:5. Iron sulfate with the concentration of 0.05mol/L is added into the solution as an inducer, iron disulfide in the high-sulfur slag is converted into elemental sulfur, and the solution is stirred, wherein the stirring speed is 100rpm, and the reaction is carried out for 120 minutes.
Step two: in the flotation stage, a proper amount of sodium hydroxide is added to adjust the pH value of the solution to control the pH value to be 2, air is blown into the solution, the air flow is 1L/min, sodium sulfide, zinc sulfate and sodium sulfite inhibitor are added, the dosage of each agent is 100 g/ton of slag, the flotation time is 10 minutes, the recovery rate of sulfur is improved by controlling the reaction conditions and adding the inhibitor, more metals such as lead, zinc and the like enter the tailing slag, and the recovery rate of sulfur is more than 95 percent.
Step three: and sending the concentrate obtained by flotation into a hot filtering section to recycle high-purity sulfur, and returning the slag after hot filtering into the slurry.
Step four: adding ammonia water into the slurry containing tailing slag and hot filter residues, controlling the concentration of the ammonia water in the slurry to be 0.5mol/L, controlling the reaction temperature to be 20 ℃, and reacting the zinc silver in the slag with the ammonia water and ammonium sulfate for 1 hour at the stirring speed of 400rpm to generate zinc ammonia complex and the silver ammonia complex to enter a liquid phase.
Step five: filtering to obtain a solution containing zinc and silver, adding formaldehyde solution into the solution in a water bath at 50 ℃ until no precipitate is generated in the solution, and filtering to recover metallic silver, wherein the recovery rate of silver is 60%.
Step six: the final zinc ammonia complex solution is used for capturing and absorbing waste gas containing carbon dioxide, the reaction temperature is controlled to be 20 ℃, the pH value is controlled to be 4.5, the zinc ammonia complex in the solution is converted into basic zinc carbonate sediment, the basic zinc carbonate product is obtained by filtering, and the recovery rate of zinc is 65%.
Example 2
Step one: 2L of oxygen pressure leaching second-stage underflow liquid is taken, the underflow liquid contains high sulfur slag, the pH value of the solution is 3, the temperature of the solution is 60 ℃, and the solid-liquid ratio (volume ratio) is 1:10. Iron sulfate with the concentration of 0.5mol/L is added into the solution as an inducer, iron disulfide in the high-sulfur slag is converted into elemental sulfur, and the solution is stirred at the stirring speed of 500rpm and reacts for 30 minutes.
Step two: in the flotation stage, a proper amount of sodium hydroxide is added to adjust the pH value of the solution, the pH value is controlled to be 4, air is blown into the solution, the air flow is 5L/min, ammonium sulfide, zinc sulfate and sodium sulfite inhibitor are added, the dosage of each agent is 300 g/ton of slag, the flotation time is 15 minutes, the recovery rate of sulfur is improved by controlling the reaction conditions and adding the inhibitor, more metals such as lead, zinc and the like enter the tailing slag, and the recovery rate of sulfur is more than 95 percent.
Step three: and sending the concentrate obtained by flotation into a hot filtering section to recycle high-purity sulfur, and returning the slag after hot filtering into the slurry.
Step four: adding ammonia water into the slurry containing tailing slag and hot filter residues, controlling the concentration of the ammonia water in the slurry to be 2mol/L, controlling the reaction temperature to be 60 ℃, and stirring at 100rpm for 3 hours, wherein zinc silver in the slag reacts with the ammonia water and ammonium sulfate to generate zinc ammonia complex and the silver ammonia complex enters a liquid phase.
Step five: filtering to obtain a solution containing zinc and silver, adding formaldehyde solution into the solution in a water bath at 70 ℃ until no precipitate is generated in the solution, and filtering to recover metallic silver, wherein the recovery rate of silver is 90%.
Step six: the final zinc ammonia complex solution is used for capturing and absorbing waste gas containing carbon dioxide, the reaction temperature is controlled to be 40 ℃, the pH value is 7, the zinc ammonia complex in the solution is converted into basic zinc carbonate sediment, the basic zinc carbonate product is obtained by filtering, and the recovery rate of zinc is 90%.
Example 3
Step one: taking 1.5L of oxygen pressure leaching second-stage underflow liquid, wherein the underflow liquid contains high sulfur slag, the pH value of the solution is 2, the temperature of the solution is 70 ℃, the solid-liquid ratio (volume ratio) is 1:7, and the solution simultaneously contains a certain amount of valuable metals such as zinc, silver and the like. Iron sulfate is added into the solution as an inducer with the concentration of 0.1 mol/L, so that iron disulfide in the high-sulfur slag is converted into elemental sulfur, and the content of the elemental sulfur is improved. And stirred at 300rpm for a reaction time of 60 minutes.
Step two: adding a proper amount of sodium hydroxide into a flotation stage to adjust the pH value of a solution, controlling the pH value to be 3, blowing air into the solution, controlling the air flow to be 3L/min, adding ammonium sulfide, sodium hydroxide, zinc sulfate and sodium sulfite inhibitor, wherein the dosage of each agent is 200 g/ton slag, the flotation time is 15 minutes, controlling the reaction conditions and adding the inhibitor, improving the recovery rate of sulfur, and enabling more metals such as lead, zinc and the like to enter the tailing slag, wherein the recovery rate of sulfur is more than 95 percent.
Step three: and sending the concentrate obtained by flotation into a hot filtering section to recycle high-purity sulfur, and returning the slag after hot filtering into the slurry.
Step four: adding ammonia water into the slurry containing tailing slag and hot filter residues, controlling the concentration of the ammonia water in the slurry to be 1mol/L, controlling the reaction temperature to be 40 ℃, stirring at 250rpm, reacting for 2 hours, and reacting zinc silver in the slag with the ammonia water and ammonium sulfate to generate zinc ammonia complex and the silver ammonia complex to enter a liquid phase.
Step five: filtering to obtain solution containing zinc and silver, adding formaldehyde or acetaldehyde solution into the solution in water bath at 60 ℃ until no precipitate is generated in the solution, and filtering to recover silver, wherein the recovery rate of silver is 75%.
Step six: the final zinc ammonia complex solution is used for capturing and absorbing waste gas containing carbon dioxide, the reaction temperature is controlled to be 30 ℃, the pH value is controlled to be 5.5, the zinc ammonia complex in the solution is converted into basic zinc carbonate sediment, the basic zinc carbonate product is obtained by filtering, and the recovery rate of zinc is 80%.
Comparative example 1
Taking 1L of oxygen pressure leaching second-stage underflow liquid, wherein the underflow liquid contains high sulfur slag, the pH value of the solution is 1, the temperature of the solution is 80 ℃, and the solid-liquid ratio (volume ratio) is 1:5. Directly sent into a flotation stage, the air flow is 1L/min, the flotation time is 10 minutes, the stirring speed is 100rpm, the flotation temperature is 80 ℃, and the sulfur recovery rate is about 75%. Concentrate is sent to the hot filtration stage. Recovery of zinc and silver is not considered.
The previous description of the disclosed embodiments is provided to enable any person skilled in the art to make or use the present invention. Various modifications to these embodiments will be readily apparent to those skilled in the art, and the generic principles defined herein may be applied to other embodiments without departing from the spirit or scope of the invention. Thus, the present invention is not intended to be limited to the embodiments shown herein but is to be accorded the widest scope consistent with the principles and novel features disclosed herein.
Claims (4)
1. A method for efficiently recycling sulfur and zinc and silver from zinc hydrometallurgy high-sulfur residues is characterized by comprising the following steps:
step one: adding an inducer into the underflow liquid of the second stage of oxygen pressure leaching, stirring, and converting iron disulfide in the high-sulfur slag into elemental sulfur;
the two-stage underflow liquid contains high sulfur slag, the pH value is 1-3, the temperature is 60-80 ℃, and the solid-liquid ratio is 1:4-1:10;
the inducer is ferric sulfate, and the concentration of the ferric sulfate in the underflow is 0.01mol/L to 0.5mol/L;
the stirring speed is 100-500 rpm, and the stirring time is 30-120 minutes;
step two: in the flotation stage, sodium hydroxide is added to adjust the pH value of the underflow to be 1.5-4, air is blown, an inhibitor is added, and flotation is carried out for 8-30 minutes;
the inhibitor is a combination of sodium sulfide, zinc sulfate and sodium sulfite, or a combination of ammonium sulfide, sodium hydroxide, zinc sulfate and sodium sulfite;
the adding mass of each medicament in the inhibitor is the same and is 100-300 g/ton of slag;
step three: the concentrate obtained by floatation is conveyed to a hot filtration section to recycle high-purity sulfur, and hot filter residues are returned to the slurry;
step four: adding ammonia water into the slurry, stirring at a certain temperature, reacting to obtain zinc ammonia complex and silver ammonia complex, and entering into a liquid phase;
step five: filtering, adding formaldehyde or acetaldehyde solution into the water bath until no precipitate is generated, and recovering silver and zinc ammonia complex solution;
step six: the zinc ammonia complex solution captures and absorbs the waste gas containing carbon dioxide, reacts under certain conditions until the solution does not generate sediment any more, and is filtered to obtain basic zinc carbonate;
the reaction temperature is 20-40 ℃ and the pH value is 4-7.
2. The method for efficiently recovering sulfur and zinc and silver from zinc hydrometallurgy high sulfur residues according to claim 1, wherein the air flow rate in the second step is 1-5L/min.
3. The method for efficiently recovering sulfur and zinc and silver from zinc hydrometallurgy high sulfur residues according to claim 1, wherein the concentration of the ammonia water in the slurry in the step four is 0.5-2 mol/L;
the temperature is 20-60 ℃, the stirring speed is 100-400 rpm, and the stirring time is 1-3 hours.
4. The method for efficiently recovering sulfur and zinc and silver from zinc hydrometallurgy high sulfur residues according to claim 1, wherein the water bath temperature in the fifth step is 50-70 ℃.
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Citations (10)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US5348713A (en) * | 1989-12-15 | 1994-09-20 | Sherritt Gordon Limited | Recovery of metal values from zinc plant residues |
| US6696037B1 (en) * | 2000-08-08 | 2004-02-24 | Dowa Mining Co., Ltd. | Method of recovering sulfur from minerals and other sulfur-containing compounds |
| JP2004292901A (en) * | 2003-03-27 | 2004-10-21 | Dowa Mining Co Ltd | Leaching method for zinc concentrate |
| BRPI0500088A (en) * | 2005-01-12 | 2006-09-05 | Assiane Clarete Adada | recycling and purification process of solid and liquid waste containing heavy metals |
| CN103773967A (en) * | 2014-02-12 | 2014-05-07 | 湘潭大学 | Method for recycling silver, copper and zinc from sintered ash in iron and steel plant |
| WO2018217083A1 (en) * | 2017-05-22 | 2018-11-29 | Elemetal Holding B.V. | Process for metal recovery by ammonia leaching and solvent extraction with gas desorption and absorption |
| CN110317957A (en) * | 2019-08-12 | 2019-10-11 | 北京矿冶科技集团有限公司 | A kind of method that zinc concentrate high-speed rail leaches |
| CN110585865A (en) * | 2019-08-27 | 2019-12-20 | 昆明理工大学 | Method for treating zinc smelting sulfur dioxide flue gas by using zinc hydrometallurgy iron-containing precipitation slag |
| CN113233426A (en) * | 2021-03-08 | 2021-08-10 | 江苏北矿金属循环利用科技有限公司 | Method for recovering sulfur from zinc oxygen pressure leaching high-sulfur slag |
| CN113976600A (en) * | 2021-10-28 | 2022-01-28 | 江苏北矿金属循环利用科技有限公司 | Harmless treatment process for toxic components of high-sulfur slag in zinc smelting |
-
2023
- 2023-08-30 CN CN202311100902.8A patent/CN116812874B/en active Active
Patent Citations (10)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US5348713A (en) * | 1989-12-15 | 1994-09-20 | Sherritt Gordon Limited | Recovery of metal values from zinc plant residues |
| US6696037B1 (en) * | 2000-08-08 | 2004-02-24 | Dowa Mining Co., Ltd. | Method of recovering sulfur from minerals and other sulfur-containing compounds |
| JP2004292901A (en) * | 2003-03-27 | 2004-10-21 | Dowa Mining Co Ltd | Leaching method for zinc concentrate |
| BRPI0500088A (en) * | 2005-01-12 | 2006-09-05 | Assiane Clarete Adada | recycling and purification process of solid and liquid waste containing heavy metals |
| CN103773967A (en) * | 2014-02-12 | 2014-05-07 | 湘潭大学 | Method for recycling silver, copper and zinc from sintered ash in iron and steel plant |
| WO2018217083A1 (en) * | 2017-05-22 | 2018-11-29 | Elemetal Holding B.V. | Process for metal recovery by ammonia leaching and solvent extraction with gas desorption and absorption |
| CN110317957A (en) * | 2019-08-12 | 2019-10-11 | 北京矿冶科技集团有限公司 | A kind of method that zinc concentrate high-speed rail leaches |
| CN110585865A (en) * | 2019-08-27 | 2019-12-20 | 昆明理工大学 | Method for treating zinc smelting sulfur dioxide flue gas by using zinc hydrometallurgy iron-containing precipitation slag |
| CN113233426A (en) * | 2021-03-08 | 2021-08-10 | 江苏北矿金属循环利用科技有限公司 | Method for recovering sulfur from zinc oxygen pressure leaching high-sulfur slag |
| CN113976600A (en) * | 2021-10-28 | 2022-01-28 | 江苏北矿金属循环利用科技有限公司 | Harmless treatment process for toxic components of high-sulfur slag in zinc smelting |
Non-Patent Citations (3)
| Title |
|---|
| Flotation separation of high sulfur lead-zinc ore;Yuan, ZT et al;《Powder Technology & Applications IV》;第454卷;第205-209页 * |
| 湿法冶炼铅银渣中回收铅银锌实验探索;何后金;白丽梅;;《云南冶金》(第06期);第23-28页 * |
| 锌湿法冶炼硫渣中硫磺化学富集工艺;张盈等;《过程工程学报》;第14卷(第1期);第57-59页 * |
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