CN111659529A - Method for separating and utilizing micro-fine particle embedded lead-zinc oxide ore by dressing and smelting - Google Patents
Method for separating and utilizing micro-fine particle embedded lead-zinc oxide ore by dressing and smelting Download PDFInfo
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- CN111659529A CN111659529A CN202010549651.1A CN202010549651A CN111659529A CN 111659529 A CN111659529 A CN 111659529A CN 202010549651 A CN202010549651 A CN 202010549651A CN 111659529 A CN111659529 A CN 111659529A
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- JQJCSZOEVBFDKO-UHFFFAOYSA-N lead zinc Chemical compound [Zn].[Pb] JQJCSZOEVBFDKO-UHFFFAOYSA-N 0.000 title claims abstract description 86
- XLOMVQKBTHCTTD-UHFFFAOYSA-N zinc oxide Inorganic materials [Zn]=O XLOMVQKBTHCTTD-UHFFFAOYSA-N 0.000 title claims abstract description 60
- 239000011787 zinc oxide Substances 0.000 title claims abstract description 60
- 238000000034 method Methods 0.000 title claims abstract description 36
- 238000003723 Smelting Methods 0.000 title claims description 10
- 239000010419 fine particle Substances 0.000 title description 2
- 238000002386 leaching Methods 0.000 claims abstract description 66
- 239000012141 concentrate Substances 0.000 claims abstract description 55
- 229910052500 inorganic mineral Inorganic materials 0.000 claims abstract description 30
- 239000011707 mineral Substances 0.000 claims abstract description 30
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims abstract description 28
- 239000003795 chemical substances by application Substances 0.000 claims abstract description 28
- 239000003112 inhibitor Substances 0.000 claims abstract description 27
- 238000000926 separation method Methods 0.000 claims abstract description 27
- 238000005188 flotation Methods 0.000 claims abstract description 22
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 claims abstract description 21
- 239000011701 zinc Substances 0.000 claims abstract description 21
- 229910052725 zinc Inorganic materials 0.000 claims abstract description 21
- AFVFQIVMOAPDHO-UHFFFAOYSA-N Methanesulfonic acid Chemical compound CS(O)(=O)=O AFVFQIVMOAPDHO-UHFFFAOYSA-N 0.000 claims abstract description 20
- 238000003756 stirring Methods 0.000 claims abstract description 15
- 239000007788 liquid Substances 0.000 claims abstract description 12
- 229940098779 methanesulfonic acid Drugs 0.000 claims abstract description 10
- 239000002002 slurry Substances 0.000 claims description 26
- 239000004088 foaming agent Substances 0.000 claims description 22
- 238000010408 sweeping Methods 0.000 claims description 22
- 229940052810 complex b Drugs 0.000 claims description 21
- QKSIFUGZHOUETI-UHFFFAOYSA-N copper;azane Chemical compound N.N.N.N.[Cu+2] QKSIFUGZHOUETI-UHFFFAOYSA-N 0.000 claims description 21
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims description 16
- 239000010949 copper Substances 0.000 claims description 16
- 229910052802 copper Inorganic materials 0.000 claims description 16
- 238000002156 mixing Methods 0.000 claims description 12
- 239000002893 slag Substances 0.000 claims description 12
- 239000007787 solid Substances 0.000 claims description 10
- CONMNFZLRNYHIQ-UHFFFAOYSA-N 3-methylbutoxymethanedithioic acid Chemical compound CC(C)CCOC(S)=S CONMNFZLRNYHIQ-UHFFFAOYSA-N 0.000 claims description 6
- 229920002134 Carboxymethyl cellulose Polymers 0.000 claims description 6
- LFQSCWFLJHTTHZ-UHFFFAOYSA-N Ethanol Chemical compound CCO LFQSCWFLJHTTHZ-UHFFFAOYSA-N 0.000 claims description 6
- 239000001768 carboxy methyl cellulose Substances 0.000 claims description 6
- 239000008112 carboxymethyl-cellulose Substances 0.000 claims description 6
- 239000003814 drug Substances 0.000 claims description 6
- 235000019353 potassium silicate Nutrition 0.000 claims description 6
- 235000019812 sodium carboxymethyl cellulose Nutrition 0.000 claims description 6
- HYHCSLBZRBJJCH-UHFFFAOYSA-M sodium hydrosulfide Chemical compound [Na+].[SH-] HYHCSLBZRBJJCH-UHFFFAOYSA-M 0.000 claims description 6
- NTHWMYGWWRZVTN-UHFFFAOYSA-N sodium silicate Chemical compound [Na+].[Na+].[O-][Si]([O-])=O NTHWMYGWWRZVTN-UHFFFAOYSA-N 0.000 claims description 6
- 229910052979 sodium sulfide Inorganic materials 0.000 claims description 6
- GRVFOGOEDUUMBP-UHFFFAOYSA-N sodium sulfide (anhydrous) Chemical compound [Na+].[Na+].[S-2] GRVFOGOEDUUMBP-UHFFFAOYSA-N 0.000 claims description 6
- 235000019832 sodium triphosphate Nutrition 0.000 claims description 6
- GGLZPLKKBSSKCX-YFKPBYRVSA-N L-ethionine Chemical compound CCSCC[C@H](N)C(O)=O GGLZPLKKBSSKCX-YFKPBYRVSA-N 0.000 claims description 5
- HQABUPZFAYXKJW-UHFFFAOYSA-O butylazanium Chemical compound CCCC[NH3+] HQABUPZFAYXKJW-UHFFFAOYSA-O 0.000 claims description 5
- HYHCSLBZRBJJCH-UHFFFAOYSA-N sodium polysulfide Chemical compound [Na+].S HYHCSLBZRBJJCH-UHFFFAOYSA-N 0.000 claims description 5
- VHUUQVKOLVNVRT-UHFFFAOYSA-N Ammonium hydroxide Chemical compound [NH4+].[OH-] VHUUQVKOLVNVRT-UHFFFAOYSA-N 0.000 claims description 3
- QPLDLSVMHZLSFG-UHFFFAOYSA-N Copper oxide Chemical compound [Cu]=O QPLDLSVMHZLSFG-UHFFFAOYSA-N 0.000 claims description 3
- 239000005751 Copper oxide Substances 0.000 claims description 3
- 239000012190 activator Substances 0.000 claims description 3
- 235000011114 ammonium hydroxide Nutrition 0.000 claims description 3
- 229910000431 copper oxide Inorganic materials 0.000 claims description 3
- 238000002425 crystallisation Methods 0.000 claims description 3
- 230000008025 crystallization Effects 0.000 claims description 3
- 238000002360 preparation method Methods 0.000 claims description 3
- CETBSQOFQKLHHZ-UHFFFAOYSA-N Diethyl disulfide Chemical compound CCSSCC CETBSQOFQKLHHZ-UHFFFAOYSA-N 0.000 claims 1
- DGAQECJNVWCQMB-PUAWFVPOSA-M Ilexoside XXIX Chemical compound C[C@@H]1CC[C@@]2(CC[C@@]3(C(=CC[C@H]4[C@]3(CC[C@@H]5[C@@]4(CC[C@@H](C5(C)C)OS(=O)(=O)[O-])C)C)[C@@H]2[C@]1(C)O)C)C(=O)O[C@H]6[C@@H]([C@H]([C@@H]([C@H](O6)CO)O)O)O.[Na+] DGAQECJNVWCQMB-PUAWFVPOSA-M 0.000 claims 1
- BESJRHHIPGWPTC-UHFFFAOYSA-N azane;copper Chemical class N.[Cu] BESJRHHIPGWPTC-UHFFFAOYSA-N 0.000 claims 1
- 229920001021 polysulfide Polymers 0.000 claims 1
- 239000005077 polysulfide Substances 0.000 claims 1
- 150000008117 polysulfides Polymers 0.000 claims 1
- 239000011734 sodium Substances 0.000 claims 1
- 229910052708 sodium Inorganic materials 0.000 claims 1
- 229910001656 zinc mineral Inorganic materials 0.000 abstract description 9
- 230000007613 environmental effect Effects 0.000 abstract description 5
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 abstract description 4
- 238000005272 metallurgy Methods 0.000 abstract description 3
- 238000000227 grinding Methods 0.000 abstract description 2
- 238000010438 heat treatment Methods 0.000 abstract description 2
- 239000011859 microparticle Substances 0.000 abstract 1
- 230000003647 oxidation Effects 0.000 abstract 1
- 238000007254 oxidation reaction Methods 0.000 abstract 1
- 238000004513 sizing Methods 0.000 abstract 1
- 238000011084 recovery Methods 0.000 description 9
- 230000002000 scavenging effect Effects 0.000 description 9
- QGZKDVFQNNGYKY-UHFFFAOYSA-N Ammonia Chemical compound N QGZKDVFQNNGYKY-UHFFFAOYSA-N 0.000 description 4
- 238000004073 vulcanization Methods 0.000 description 4
- LSNNMFCWUKXFEE-UHFFFAOYSA-M Bisulfite Chemical compound OS([O-])=O LSNNMFCWUKXFEE-UHFFFAOYSA-M 0.000 description 3
- 239000005083 Zinc sulfide Substances 0.000 description 3
- 239000002585 base Substances 0.000 description 3
- 239000012535 impurity Substances 0.000 description 3
- 229910000464 lead oxide Inorganic materials 0.000 description 3
- YEXPOXQUZXUXJW-UHFFFAOYSA-N oxolead Chemical compound [Pb]=O YEXPOXQUZXUXJW-UHFFFAOYSA-N 0.000 description 3
- 229910052984 zinc sulfide Inorganic materials 0.000 description 3
- 239000002253 acid Substances 0.000 description 2
- 239000003513 alkali Substances 0.000 description 2
- WUOACPNHFRMFPN-UHFFFAOYSA-N alpha-terpineol Chemical group CC1=CCC(C(C)(C)O)CC1 WUOACPNHFRMFPN-UHFFFAOYSA-N 0.000 description 2
- 229910021529 ammonia Inorganic materials 0.000 description 2
- 150000003863 ammonium salts Chemical class 0.000 description 2
- SQIFACVGCPWBQZ-UHFFFAOYSA-N delta-terpineol Natural products CC(C)(O)C1CCC(=C)CC1 SQIFACVGCPWBQZ-UHFFFAOYSA-N 0.000 description 2
- 229910052751 metal Inorganic materials 0.000 description 2
- 239000002184 metal Substances 0.000 description 2
- 125000002496 methyl group Chemical group [H]C([H])([H])* 0.000 description 2
- 150000003839 salts Chemical class 0.000 description 2
- 229940116411 terpineol Drugs 0.000 description 2
- 229910021532 Calcite Inorganic materials 0.000 description 1
- OYPRJOBELJOOCE-UHFFFAOYSA-N Calcium Chemical compound [Ca] OYPRJOBELJOOCE-UHFFFAOYSA-N 0.000 description 1
- FYYHWMGAXLPEAU-UHFFFAOYSA-N Magnesium Chemical compound [Mg] FYYHWMGAXLPEAU-UHFFFAOYSA-N 0.000 description 1
- 238000009825 accumulation Methods 0.000 description 1
- 230000004913 activation Effects 0.000 description 1
- 239000000654 additive Substances 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 239000011575 calcium Substances 0.000 description 1
- 229910052791 calcium Inorganic materials 0.000 description 1
- 150000003841 chloride salts Chemical class 0.000 description 1
- 150000001875 compounds Chemical class 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000005363 electrowinning Methods 0.000 description 1
- 230000002708 enhancing effect Effects 0.000 description 1
- 230000002209 hydrophobic effect Effects 0.000 description 1
- 230000005661 hydrophobic surface Effects 0.000 description 1
- 150000002500 ions Chemical class 0.000 description 1
- -1 lead-zinc metals Chemical class 0.000 description 1
- 239000011777 magnesium Substances 0.000 description 1
- 229910052749 magnesium Inorganic materials 0.000 description 1
- 150000002739 metals Chemical class 0.000 description 1
- 238000005065 mining Methods 0.000 description 1
- 238000005456 ore beneficiation Methods 0.000 description 1
- 238000004321 preservation Methods 0.000 description 1
- 239000010453 quartz Substances 0.000 description 1
- 239000002994 raw material Substances 0.000 description 1
- 230000009257 reactivity Effects 0.000 description 1
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N silicon dioxide Inorganic materials O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 description 1
- 238000000638 solvent extraction Methods 0.000 description 1
- 238000001179 sorption measurement Methods 0.000 description 1
- 229910052569 sulfide mineral Inorganic materials 0.000 description 1
- 239000002699 waste material Substances 0.000 description 1
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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
- B03B9/00—General arrangement of separating plant, e.g. flow sheets
-
- 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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- 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
- B03D1/001—Flotation agents
- B03D1/002—Inorganic compounds
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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
- B03D1/001—Flotation agents
- B03D1/018—Mixtures of inorganic and organic compounds
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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
- B03D2201/00—Specified effects produced by the flotation agents
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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
- B03D2201/00—Specified effects produced by the flotation agents
- B03D2201/02—Collectors
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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
- B03D2201/00—Specified effects produced by the flotation agents
- B03D2201/06—Depressants
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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
- B03D2203/00—Specified materials treated by the flotation agents; Specified applications
- B03D2203/02—Ores
- B03D2203/04—Non-sulfide ores
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- Chemical & Material Sciences (AREA)
- Inorganic Chemistry (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
本发明公开了一种微细粒嵌布型氧化铅锌矿选冶分离与利用的方法,属于矿物加工技术领域。本发明将微细粒嵌布型氧化铅锌矿破碎、磨矿、调浆后,添加组合抑制剂协同抑制脉石矿物,添加铜氨络合物梯级活化铅锌矿物,添加组合硫化剂协同硫化氧化铅锌矿物,添加组合捕收剂使硫化后的氧化铅锌矿物强化疏水,有效地抛除矿石中的脉石矿物,得到铅锌浮选混合精矿,然后采用硫酸浸出其中的锌矿物,固液分离后得到含铅浸出渣和含锌浸出液;含铅浸出渣采用甲基磺酸为浸出剂,通过加温搅拌浸出铅矿物,得到含铅浸出液;本发明经济高效地解决了微细粒嵌布型氧化铅锌矿难以分离的技术难题,提高了铅锌资源的综合利用率,社会、环境和经济效益显著。
The invention discloses a method for beneficiation, metallurgy, separation and utilization of fine-grain embedded lead-zinc oxide ore, belonging to the technical field of mineral processing. In the present invention, after crushing, grinding and sizing of fine-grained embedded lead-zinc oxide ore, a combination inhibitor is added to synergistically suppress gangue minerals, a copper-ammine complex is added to activate lead-zinc minerals in a step-by-step manner, and a combined vulcanizing agent is added to synergize sulfide oxidation. Lead-zinc minerals, adding a combined collector to strengthen the hydrophobicity of the sulfided lead-zinc oxide minerals, effectively remove the gangue minerals in the ore, and obtain the lead-zinc flotation mixed concentrate, and then use sulfuric acid to leaching the zinc minerals in it, solidify the ore. The lead-containing leaching residue and the zinc-containing leaching solution are obtained after liquid separation; the lead-containing leaching residue adopts methanesulfonic acid as the leaching agent, and the lead minerals are leached by heating and stirring to obtain the lead-containing leaching solution; the invention solves the problem of micro-particle embedding economically and efficiently. The technical problem that the lead-zinc oxide ore is difficult to separate has improved the comprehensive utilization rate of lead-zinc resources, and the social, environmental and economic benefits are remarkable.
Description
技术领域technical field
本发明涉及一种微细粒嵌布型氧化铅锌矿选冶分离与利用的方法,属于矿物加工技术领域。The invention relates to a method for beneficiation, metallurgy, separation and utilization of fine-grained embedded lead-zinc oxide ore, belonging to the technical field of mineral processing.
背景技术Background technique
长期以来,铅锌冶炼一直以硫化矿物为原料,然而,随着有限的硫化铅锌资源不断消耗,现存的硫化铅锌矿物已不能满足日益增长的金属需求,因此,为了确保铅锌金属的保有量,降低对外依存度,氧化铅锌矿的高效开发利用已成必然。氧化铅锌矿是我国重要的铅锌矿产资源,储量丰富,但具有贫、细、杂、泥化严重等特点,且这类矿石含有大量的可溶性盐;与硫化铅锌矿物相比,氧化铅锌矿物具有更高的溶解度和表面亲水性,因此大量的氧化铅锌矿产资源至今未被充分开发利用,且相当一部分低品位氧化铅锌矿开采后堆存于矿山附近,不仅造成资源浪费,还会引起环境问题。For a long time, lead-zinc smelting has always used sulfide minerals as raw materials. However, with the continuous consumption of limited lead-zinc sulfide resources, the existing lead-zinc sulfide minerals can no longer meet the growing demand for metals. Therefore, in order to ensure the preservation of lead-zinc metals The high-efficiency development and utilization of lead-zinc oxide ore has become inevitable. Lead-zinc oxide ore is an important lead-zinc mineral resource in my country, with abundant reserves, but it has the characteristics of being poor, fine, miscellaneous, and seriously muddy, and this type of ores contains a large amount of soluble salts; compared with lead-zinc sulfide minerals, lead oxide Zinc minerals have higher solubility and surface hydrophilicity, so a large number of lead-zinc oxide mineral resources have not been fully exploited yet, and a considerable part of low-grade lead-zinc oxide minerals are stored near the mine after mining, which not only causes resource waste, It can also cause environmental problems.
选矿是富集氧化铅矿的重要方法,但由于氧化铅锌矿中矿物种类较多,矿石结构复杂,伴生组分不稳定,含有大量活化的石英、方解石以及可溶性盐类矿物,因此氧化铅锌矿通过选矿的方法进行回收仍存在一些问题。火法工艺能够处理铅锌品位较高,钙镁含量较低的氧化铅锌矿,当氧化铅锌矿中钙镁含量较高时,火法工艺存在大量的问题,如物料碱度大、黏度大,铅锌损失高,添加剂用量大;另外,火法回收铅锌存在不同程度的铅蒸汽污染,环境风险较大。目前,浸出工艺常用来提取氧化铅锌矿,硫酸浸出后得到的浸出液可通过溶剂萃取或电积工艺直接提取金属锌,当矿石中钙镁含量较高时,酸耗较高、浸出液中杂质含量较高,后续处理成本较高;采用碱性体系、铵盐体系、氯盐体系浸出氧化铅锌矿时,浸出渣量大,渣中碱含量高,同时会形成难以处理的强碱浸渣,大量的强碱浸渣堆存会引起二次污染,环境风险较大,这极大地限制了该工艺的推广。Ore beneficiation is an important method for enriching lead oxide ore. However, due to the large number of minerals in lead oxide ore, the ore structure is complex, the associated components are unstable, and it contains a large amount of activated quartz, calcite and soluble salt minerals. There are still some problems in the recovery of ore by beneficiation. The pyrotechnic process can process lead-zinc oxide ore with high grade of lead and zinc and low calcium and magnesium content. Large, lead and zinc losses are high, and the amount of additives is large; in addition, the pyrotechnic recovery of lead and zinc has different degrees of lead vapor pollution, and the environmental risk is relatively large. At present, the leaching process is commonly used to extract lead-zinc oxide ore, and the leaching solution obtained after sulfuric acid leaching can directly extract metallic zinc through solvent extraction or electrowinning process. higher, and the cost of follow-up treatment is higher; when using alkaline system, ammonium salt system, and chloride salt system to leaching lead-zinc oxide ore, the amount of leaching slag is large, the alkali content in the slag is high, and at the same time, a strong alkali leaching slag that is difficult to handle will be formed. The accumulation of a large amount of strong alkaline leaching residue will cause secondary pollution and cause great environmental risks, which greatly limits the promotion of this process.
微细粒嵌布型氧化铅锌矿是一类极难分离的氧化铅锌矿,即使采用细磨工艺仍然不能将矿石中的氧化铅锌矿物有效解离,采用传统的选矿方法铅锌分离困难,产品中铅锌互含率高,难以满足后续处理工艺的需要。因此亟需研发一种能够高效分离与利用微细粒嵌布型氧化铅锌矿的新工艺,最大程度提高该类资源的综合利用率,缓解铅锌供需矛盾,增加企业效益。Fine-grained embedded lead-zinc oxide ore is a kind of lead-zinc oxide ore that is extremely difficult to separate. Even if the fine grinding process is used, the lead-zinc oxide minerals in the ore cannot be effectively dissociated. It is difficult to separate lead and zinc by traditional beneficiation methods. The mutual content of lead and zinc in the product is high, and it is difficult to meet the needs of the subsequent treatment process. Therefore, it is urgent to develop a new process that can efficiently separate and utilize fine-grain embedded lead-zinc oxide ore, so as to maximize the comprehensive utilization rate of such resources, alleviate the contradiction between supply and demand of lead and zinc, and increase enterprise benefits.
发明内容SUMMARY OF THE INVENTION
本发明针对微细粒嵌布型氧化铅锌矿采用单一浮选工艺分离困难,采用单一浸出方法成本高、浸出液中杂质元素高的技术难题,提供一种微细粒嵌布型氧化铅锌矿选冶分离与利用的方法,即将微细粒嵌布型氧化铅锌矿破碎、磨矿、调浆后,添加组合抑制剂协同抑制脉石矿物,添加铜氨络合物梯级活化铅锌矿物,添加组合硫化剂协同硫化氧化铅锌矿物,添加组合捕收剂使硫化后的氧化铅锌矿物强化疏水,有效地抛除矿石中的脉石矿物,得到铅锌浮选混合精矿,然后采用硫酸浸出其中的锌矿物,固液分离后得到含铅浸出渣和含锌浸出液;含铅浸出渣采用甲基磺酸为浸出剂,通过加温搅拌浸出铅矿物,得到含铅浸出液;本发明经济高效地解决了微细粒嵌布型氧化铅锌矿难以分离的技术难题,提高了铅锌资源的综合利用率,社会、环境和经济效益显著。The invention provides a fine-grain embedded type lead-zinc oxide ore dressing and smelting in order to solve the technical problems that the fine-grain embedded type lead-zinc oxide ore is difficult to separate by a single flotation process, the single leaching method is high in cost, and the impurity elements in the leaching solution are high. The method of separation and utilization is that after the fine-grain embedded lead-zinc oxide ore is crushed, ground, and pulped, a combination inhibitor is added to synergistically inhibit the gangue minerals, a copper-ammine complex is added to activate the lead-zinc minerals, and a combined sulfide is added. The sulfided lead-zinc oxide minerals are sulfided together with the combined collectors, and the sulfided lead-zinc oxide minerals are strengthened and hydrophobic, and the gangue minerals in the ore are effectively thrown away to obtain the lead-zinc flotation mixed concentrate, and then sulfuric acid is used to leaching the ore in it. Zinc minerals, lead-containing leaching residue and zinc-containing leaching solution are obtained after solid-liquid separation; lead-containing leaching residues use methanesulfonic acid as leaching agent, and lead minerals are leached by heating and stirring to obtain lead-containing leaching solution; The technical problem that the fine-grained embedded lead-zinc oxide ore is difficult to separate has improved the comprehensive utilization rate of lead-zinc resources, and has significant social, environmental and economic benefits.
一种微细粒嵌布型氧化铅锌矿选冶分离与利用的方法,具体步骤如下:A method for beneficiation, metallurgy, separation and utilization of fine-grain embedded lead-zinc oxide ore, the specific steps are as follows:
(1)将微细粒嵌布型氧化铅锌矿破碎、磨细至-74μm粒级的质量百分含量占80~90%,调浆至矿浆质量百分浓度为30~40%,在矿浆中依次加入组合抑制剂、铜氨络合物A、组合硫化剂、铜氨络合物B、组合捕收剂和起泡剂,进行粗选作业得到粗选精矿和粗选尾矿;(1) The mass percentage of fine-grained embedded lead-zinc oxide ore is crushed and ground to -74μm size, accounting for 80-90%, and the mass percentage of the pulp is adjusted to 30-40%. The combined inhibitor, the cupro-ammine complex A, the combined vulcanizing agent, the cupro-ammine complex B, the combined collector and the foaming agent are added in sequence, and the roughing operation is carried out to obtain roughing concentrate and roughing tailings;
(2)在步骤(1)粗选尾矿中依次加入组合抑制剂、铜氨络合物A、组合硫化剂、铜氨络合物B、组合捕收剂和起泡剂,进行一次扫选作业得到一次扫选精矿和一次扫选尾矿,其中一次扫选精矿返回调浆并入粗选作业;在一次扫选尾矿中依次加入组合抑制剂、铜氨络合物A、组合硫化剂、铜氨络合物B、组合捕收剂和起泡剂,进行二次扫选作业得到二次扫选精矿和浮选尾矿,其中二次扫选精矿返回调浆并入一次扫选作业;(2) in step (1) roughing tailings are sequentially added combined inhibitor, cuproammine complex A, combined vulcanizing agent, cuproammine complex B, combined collector and foaming agent to carry out a sweep The operation obtains a sweeping concentrate and a sweeping tailings, of which the first sweeping concentrate is returned to the slurry to be incorporated into the roughing operation; in the first sweeping tailings, a combination inhibitor, a copper ammonia complex A, a combination of Sulfurizing agent, copper ammonia complex B, combined collector and foaming agent, carry out secondary sweeping operation to obtain secondary sweeping concentrate and flotation tailings, of which the secondary sweeping concentrate is returned to slurry mixing and incorporated One scan job;
(3)将步骤(1)粗选精矿进行一次精选作业得到一次精选精矿和一次精选尾矿,一次精选尾矿返回调浆并入粗选作业,一次精选精矿进行二次精选作业得到铅锌浮选混合精矿和二次精选尾矿,其中二次精选尾矿返回调浆并入一次精选作业;(3) Perform a beneficiation operation on the rougher concentrate in step (1) to obtain a beneficiary concentrate and a beneficiary tailings, a beneficiation tailings are returned to the slurry to be incorporated into the roughing operation, and a beneficiation concentrate is carried out The secondary beneficiation operation obtains lead-zinc flotation mixed concentrate and secondary beneficiation tailings, of which the secondary beneficiary tailings are returned to slurry mixing and merged into the primary beneficiation operation;
(4)将步骤(3)所得的铅锌浮选混合精矿调浆,在温度为45~75℃的搅拌条件下,进行硫酸浸出60~120min,固液分离后得到含铅浸出渣和含锌浸出液;(4) Slurrying the lead-zinc flotation mixed concentrate obtained in step (3), and performing sulfuric acid leaching for 60-120 min under stirring conditions at a temperature of 45-75°C, and obtaining lead-containing leaching slag and lead-containing leaching slag after solid-liquid separation Zinc leaching solution;
(5)将步骤(4)所得的含铅浸出渣调浆,在温度为40~60℃的搅拌条件下,进行甲基磺酸浸出35~55min,固液分离后得到含铅溶液和浸出尾渣。(5) slurrying the lead-containing leaching residue obtained in step (4), under stirring conditions at a temperature of 40 to 60° C., leaching with methanesulfonic acid for 35 to 55 minutes, and obtaining a lead-containing solution and a leaching tail after solid-liquid separation scum.
以每吨微细粒嵌布型氧化铅锌矿计,所述步骤(1)矿浆中加入组合抑制剂1400~2200g、铜氨络合物A 500~900g、组合硫化剂3200~7600g、铜氨络合物B300~600g、组合捕收剂800~1400g、起泡剂60~80g。Calculated per ton of fine-grained embedded lead-zinc oxide ore, in the step (1), add 1400-2200 g of combined inhibitor, 500-900 g of copper ammonia complex A, 3200 to 7600 g of combined vulcanizing agent, and copper ammonia complex Compound B 300-600g, combined collector 800-1400g, foaming agent 60-80g.
以每吨微细粒嵌布型氧化铅锌矿计,所述步骤(2)粗选尾矿中依次加入组合抑制剂700~1100g、铜氨络合物A 250~450g、组合硫化剂1600~3800g、铜氨络合物B150~300g、组合捕收剂400~700g、起泡剂30~40g。Calculated per ton of fine-grained embedded lead-zinc oxide ore, the step (2) is followed by adding 700-1100 g of combined inhibitor, 250-450 g of copper ammine complex A, and 1600-3800 g of combined vulcanizing agent into the roughing tailings in step (2). , Copper ammonia complex B150~300g, combined collector 400~700g, foaming agent 30~40g.
以每吨微细粒嵌布型氧化铅锌矿计,所述步骤(2)一次扫选尾矿中依次加入组合抑制剂350~550g、铜氨络合物A 125~225g、组合硫化剂800~1900g、铜氨络合物B75~150g、组合捕收剂200~350g、起泡剂15~20g。In the step (2), 350-550 g of combined inhibitor, 125-225 g of copper-ammine complex A, and 800-800 g of combined vulcanizing agent are sequentially added to the tailings in step (2) in terms of per ton of fine-grained embedded lead-zinc oxide ore. 1900g, copper ammonia complex B75-150g, combined collector 200-350g, foaming agent 15-20g.
所述步骤(4)铅锌浮选混合精矿调浆浓度为矿浆的液固质量比为3:1,硫酸浓度为55~135g/L。In the step (4) of the lead-zinc flotation mixed concentrate, the slurry concentration is such that the liquid-solid mass ratio of the slurry is 3:1, and the sulfuric acid concentration is 55-135 g/L.
所述步骤(5)含铅浸出渣调浆浓度为矿浆的液固质量比为3:1,甲基磺酸的浓度为45~95g/L。In the step (5), the concentration of the lead-containing leaching slag is adjusted so that the liquid-solid mass ratio of the pulp is 3:1, and the concentration of the methanesulfonic acid is 45-95 g/L.
所述组合抑制剂包括水玻璃、三聚磷酸钠和羧甲基纤维素,其中水玻璃、三聚磷酸钠和羧甲基纤维素的质量比为4:3:3。The combined inhibitor includes water glass, sodium tripolyphosphate and carboxymethyl cellulose, wherein the mass ratio of water glass, sodium tripolyphosphate and carboxymethyl cellulose is 4:3:3.
所述组合硫化剂包括硫化钠、多硫化钠和硫氢化钠,其中硫化钠、多硫化钠和硫氢化钠的质量比为5:2:3。The combined vulcanizing agent includes sodium sulfide, sodium polysulfide and sodium hydrosulfide, wherein the mass ratio of sodium sulfide, sodium polysulfide and sodium hydrosulfide is 5:2:3.
所述组合捕收剂包括异戊基黄药、乙硫氮和丁铵黑药,其中异戊基黄药、乙硫氮和丁铵黑药的质量比为4:3:3。The combined collector includes isopentyl xanthate, ethionine and butylammonium black medicine, wherein the mass ratio of isopentyl xanthate, ethionine and butylammonium black medicine is 4:3:3.
所述铜氨络合物A和铜氨络合物B为相同的铜氨络合物,其制备方法为:Described cuprammonium complex A and cuprammonium complex B are identical cuprammonium complexes, and the preparation method thereof is:
1)采用浓氨水对高纯度氧化铜矿物进行搅拌浸出得到铜氨络合物溶液;1) using concentrated ammonia water to carry out stirring and leaching of high-purity copper oxide minerals to obtain a copper ammine complex solution;
2)将步骤1)铜氨络合物溶液置于乙醇溶液中进行多次结晶,得到活化剂铜氨络合物。2) The copper ammine complex solution of step 1) is placed in an ethanol solution for multiple crystallization to obtain the activator copper ammine complex.
优选的,所述起泡剂为松醇油。Preferably, the foaming agent is terpineol oil.
本发明的有益效果:Beneficial effects of the present invention:
(1)本发明采用铜氨络合物对微细粒嵌布型氧化铅锌矿中的铅锌矿物进行梯级活化,增强矿物表面反应活性,促进矿物表面硫化反应,形成更厚更稳定的硫化层,实现强化硫化;同时铜氨络离子还能选择性地吸附在强化硫化后的氧化铅锌矿物表面,从而增强捕收剂吸附,提高矿物表面疏水性;(1) the present invention adopts the copper ammine complex to carry out the cascade activation of the lead-zinc minerals in the fine-grained embedded lead-zinc oxide ore, enhances the mineral surface reactivity, promotes the mineral surface vulcanization reaction, and forms a thicker and more stable sulfide layer , to achieve enhanced vulcanization; at the same time, copper ammonia complex ions can also selectively adsorb on the surface of lead-zinc oxide minerals after enhanced vulcanization, thereby enhancing the adsorption of collectors and improving the hydrophobicity of the mineral surface;
(2)本发明通过组合抑制剂协同抑制脉石矿物,铜氨络合物梯级活化氧化铅锌矿物,组合硫化剂协同硫化氧化铅锌矿物,添加组合捕收剂使硫化后的矿物表面强化疏水,不仅实现了矿石中铅锌矿物的有效富集,而且抛除大量的不利于浸出工艺的脉石矿物,大幅减少了浸出过程中的矿石量,降低了浸出液中杂质元素的含量,降低了矿石处理成本;(2) The present invention synergistically inhibits gangue minerals by combining inhibitors, activates lead-zinc oxide minerals in a step-by-step manner with copper ammine complexes, combines vulcanizing agents to synergistically sulfide lead-zinc oxide minerals, and adds combined collectors to strengthen the hydrophobic surface of the vulcanized minerals. , which not only realizes the effective enrichment of lead and zinc minerals in the ore, but also removes a large number of gangue minerals that are not conducive to the leaching process, greatly reducing the amount of ore in the leaching process, reducing the content of impurity elements in the leaching solution, and reducing the amount of ore. processing costs;
(3)本发明采用硫酸浸出铅锌浮选混合精矿,得到含铅浸出渣和含锌浸出液,含铅浸出渣采用甲基磺酸浸出,得到含铅浸出液和浸出尾渣,实现了微细粒嵌布型氧化铅锌矿的高效回收和分离,经济高效地解决了微细粒嵌布型氧化铅锌矿采用单一浮选工艺分离困难,采用单一浸出方法成本高、浸出液中杂质元素高的技术难题。(3) the present invention adopts sulfuric acid leaching lead-zinc flotation mixed concentrate to obtain lead-containing leaching residue and zinc-containing leaching solution, and the lead-containing leaching residue is leached with methanesulfonic acid to obtain lead-containing leaching solution and leaching tailings, and realizes fine particles The efficient recovery and separation of embedded lead-zinc oxide ore economically and efficiently solves the technical problems that the fine-grain embedded lead-zinc oxide ore is difficult to separate by a single flotation process, the single leaching method is costly, and the impurity elements in the leaching solution are high. .
附图说明Description of drawings
图1为本发明的工艺流程图。Fig. 1 is a process flow diagram of the present invention.
具体实施方式Detailed ways
下面结合具体实施例对本发明作进一步详细说明,但本发明的保护范围并不限于所述内容。The present invention will be described in further detail below with reference to specific embodiments, but the protection scope of the present invention is not limited to the content.
本发明以下实施例中组合抑制剂包括水玻璃、三聚磷酸钠和羧甲基纤维素,其中水玻璃、三聚磷酸钠和羧甲基纤维素的质量比为4:3:3;组合硫化剂包括硫化钠、多硫化钠和硫氢化钠,其中硫化钠、多硫化钠和硫氢化钠的质量比为5:2:3;组合捕收剂包括异戊基黄药、乙硫氮和丁铵黑药,其中异戊基黄药、乙硫氮和丁铵黑药的质量比为4:3:3;起泡剂为松醇油;In the following embodiments of the present invention, the combined inhibitor includes water glass, sodium tripolyphosphate and carboxymethyl cellulose, wherein the mass ratio of water glass, sodium tripolyphosphate and carboxymethyl cellulose is 4:3:3; the combined vulcanization Agents include sodium sulfide, sodium polysulfide and sodium hydrosulfide, wherein the mass ratio of sodium sulfide, sodium polysulfide and sodium hydrosulfide is 5:2:3; combined collectors include isopentyl xanthate, ethionine and butylated Ammonium black medicine, wherein the mass ratio of isopentyl xanthate, ethionine and butylammonium black medicine is 4:3:3; foaming agent is terpineol oil;
铜氨络合物A和铜氨络合物B为相同的铜氨络合物,其制备方法为:Cuprammonium complex A and cuprammonium complex B are identical cuprammonium complexes, and their preparation method is:
1)采用浓氨水对高纯度氧化铜矿物进行搅拌浸出得到铜氨络合物溶液;1) using concentrated ammonia water to carry out stirring and leaching of high-purity copper oxide minerals to obtain a copper ammine complex solution;
2)将步骤1)铜氨络合物溶液置于乙醇溶液中进行多次结晶,得到活化剂铜氨络合物。2) The copper ammine complex solution of step 1) is placed in an ethanol solution for multiple crystallization to obtain the activator copper ammine complex.
实施例1:本实施例的微细粒嵌布型氧化铅锌矿中铅的质量百分数含量为4.3%,锌的质量百分数含量为5.4%;Embodiment 1: The mass percentage content of lead in the fine-grained embedded lead-zinc oxide ore of this embodiment is 4.3%, and the mass percentage content of zinc is 5.4%;
如图1所示,一种微细粒嵌布型氧化铅锌矿选冶分离与利用的方法,具体步骤如下:As shown in Figure 1, a kind of method for beneficiation and smelting separation and utilization of fine-grain embedded type lead-zinc oxide ore, the concrete steps are as follows:
(1)将微细粒嵌布型氧化铅锌矿破碎、磨细至-74μm粒级的质量百分含量占90%,调浆至矿浆质量百分浓度为30%,以每吨微细粒嵌布型氧化铅锌矿计,在矿浆中依次加入1400g组合抑制剂、500g铜氨络合物A、3200g组合硫化剂、300g铜氨络合物B、800g组合捕收剂和60g起泡剂,进行粗选作业得到粗选精矿和粗选尾矿;(1) The fine-grained embedded lead-zinc oxide ore is crushed and ground to -74μm with a mass percentage of 90%, and the pulp is slurried until the mass percentage concentration of the pulp is 30%. Type lead-zinc oxide ore, add 1400g combined inhibitor, 500g cupro-ammine complex A, 3200g combined vulcanizing agent, 300g cupro-ammine complex B, 800g combined collector and 60g foaming agent in the pulp in sequence, and carry out The roughing operation produces rougher concentrate and rougher tailings;
(2)以每吨微细粒嵌布型氧化铅锌矿计,在步骤(1)粗选尾矿中依次加入700g组合抑制剂、250g铜氨络合物A、1600g组合硫化剂、150g铜氨络合物B、400g组合捕收剂和30g起泡剂,进行一次扫选作业得到一次扫选精矿和一次扫选尾矿,其中一次扫选精矿返回调浆并入粗选作业;在一次扫选尾矿中依次加入350g组合抑制剂、125g铜氨络合物A、800g组合硫化剂、75g铜氨络合物B、200g组合捕收剂和15g起泡剂,进行二次扫选作业得到二次扫选精矿和浮选尾矿,其中二次扫选精矿返回调浆并入一次扫选作业;(2) in terms of each ton of fine-grain embedded lead-zinc oxide ore, in step (1) roughing tailings, add 700g combined inhibitor, 250g cupro ammonia complex A, 1600g combined vulcanizing agent, 150g cupro ammonia Complex B, 400g combined collector and 30g foaming agent, carry out a sweep operation to obtain a sweep concentrate and a sweep tailing, of which the first sweep concentrate is returned to slurry mixing and incorporated into the roughing operation; 350g of combined inhibitor, 125g of cupro-ammine complex A, 800g of combined vulcanizing agent, 75g of cupro-ammine complex B, 200g of combined collector and 15g of foaming agent were added to the tailings in the first sweep, and the secondary sweep was carried out. The operation obtains secondary scavenging concentrates and flotation tailings, of which the secondary scavenging concentrates are returned to slurry and merged into the primary scavenging operation;
(3)将步骤(1)粗选精矿进行一次精选作业得到一次精选精矿和一次精选尾矿,一次精选尾矿返回调浆并入粗选作业,一次精选精矿进行二次精选作业得到铅锌浮选混合精矿和二次精选尾矿,其中二次精选尾矿返回调浆并入一次精选作业;(3) Perform a beneficiation operation on the rougher concentrate in step (1) to obtain a beneficiary concentrate and a beneficiary tailings, a beneficiation tailings are returned to the slurry to be incorporated into the roughing operation, and a beneficiation concentrate is carried out The secondary beneficiation operation obtains lead-zinc flotation mixed concentrate and secondary beneficiation tailings, of which the secondary beneficiary tailings are returned to slurry mixing and merged into the primary beneficiation operation;
(4)将步骤(3)所得的铅锌浮选混合精矿调浆至矿浆的液固质量比为3:1,在温度为45℃、硫酸浓度为55g/L的搅拌条件下,进行硫酸浸出120min,固液分离后得到含铅浸出渣和含锌浸出液;(4) the lead-zinc flotation mixed concentrate obtained in step (3) is slurried to a liquid-solid mass ratio of 3:1, and the sulfuric acid is carried out under the stirring condition that the temperature is 45° C. and the sulfuric acid concentration is 55 g/L. After leaching for 120min, lead-containing leaching residue and zinc-containing leaching solution are obtained after solid-liquid separation;
(5)将步骤(4)所得的含铅浸出渣调浆至矿浆的液固质量比为3:1,在温度为40℃、甲基磺酸浓度为45g/L的搅拌条件下,进行甲基磺酸浸出55min,固液分离后得到含铅溶液和浸出尾渣;(5) the liquid-solid mass ratio of the lead-containing leaching slag of step (4) gained to slurry is 3:1, under the stirring condition that temperature is 40 ℃, methanesulfonic acid concentration is 45g/L, carry out methyl acid The base sulfonic acid was leached for 55 minutes, and the lead-containing solution and the leaching tailings were obtained after solid-liquid separation;
本实施例中铅的回收率为87.8%,锌的回收率为88.3%。In this example, the recovery rate of lead was 87.8%, and the recovery rate of zinc was 88.3%.
实施例2:本实施例的微细粒嵌布型氧化铅锌矿中铅的质量百分数含量为5.4%,锌的质量百分数含量为6.8%;Embodiment 2: the mass percentage content of lead in the fine-grained embedded lead-zinc oxide ore of this embodiment is 5.4%, and the mass percentage content of zinc is 6.8%;
如图1所示,一种微细粒嵌布型氧化铅锌矿选冶分离与利用的方法,具体步骤如下:As shown in Figure 1, a kind of method for beneficiation and smelting separation and utilization of fine-grain embedded type lead-zinc oxide ore, the concrete steps are as follows:
(1)将微细粒嵌布型氧化铅锌矿破碎、磨细至-74μm粒级的质量百分含量占85%,调浆至矿浆质量百分浓度为35%,以每吨微细粒嵌布型氧化铅锌矿计,在矿浆中依次加入1800g组合抑制剂、700g铜氨络合物A、5400g组合硫化剂、450g铜氨络合物B、1100g组合捕收剂和70g起泡剂,进行粗选作业得到粗选精矿和粗选尾矿;(1) The fine-grained embedded lead-zinc oxide ore is crushed and ground to -74μm with a mass percentage of 85%, and the pulp is slurried until the mass percentage concentration of the ore pulp is 35%. Type lead-zinc oxide ore, add 1800g combined inhibitor, 700g cupro-ammine complex A, 5400g combined vulcanizing agent, 450g cupro-ammine complex B, 1100g combined collector and 70g foaming agent in the pulp in sequence, and carry out The roughing operation produces rougher concentrate and rougher tailings;
(2)以每吨微细粒嵌布型氧化铅锌矿计,在步骤(1)粗选尾矿中依次加入900g组合抑制剂、350g铜氨络合物A、2700g组合硫化剂、225g铜氨络合物B、550g组合捕收剂和35g起泡剂,进行一次扫选作业得到一次扫选精矿和一次扫选尾矿,其中一次扫选精矿返回调浆并入粗选作业;在一次扫选尾矿中依次加入450g组合抑制剂、175g铜氨络合物A、1350g组合硫化剂、112g铜氨络合物B、275g组合捕收剂和17g起泡剂,进行二次扫选作业得到二次扫选精矿和浮选尾矿,其中二次扫选精矿返回调浆并入一次扫选作业;(2) In terms of each ton of fine-grained embedded type lead-zinc oxide ore, in step (1) the roughing tailings were successively added with 900g combined inhibitor, 350g copper ammonia complex A, 2700g combined vulcanizing agent, 225g copper ammonia Complex B, 550g combined collector and 35g foaming agent, carry out a sweeping operation to obtain a sweeping concentrate and a sweeping tailings, of which the first sweeping concentrate is returned to slurry mixing and is incorporated into the roughing operation; 450g of combined inhibitor, 175g of copper ammonia complex A, 1350g of combined vulcanizing agent, 112g of copper ammonia complex B, 275g of combined collector and 17g of foaming agent were added to the tailings in the first sweep, and the second sweep was carried out. The operation obtains secondary scavenging concentrates and flotation tailings, of which the secondary scavenging concentrates are returned to slurry and merged into the primary scavenging operation;
(3)将步骤(1)粗选精矿进行一次精选作业得到一次精选精矿和一次精选尾矿,一次精选尾矿返回调浆并入粗选作业,一次精选精矿进行二次精选作业得到铅锌浮选混合精矿和二次精选尾矿,其中二次精选尾矿返回调浆并入一次精选作业;(3) Perform a beneficiation operation on the rougher concentrate in step (1) to obtain a beneficiary concentrate and a beneficiary tailings, a beneficiation tailings are returned to the slurry to be incorporated into the roughing operation, and a beneficiation concentrate is carried out The secondary beneficiation operation obtains lead-zinc flotation mixed concentrate and secondary beneficiation tailings, of which the secondary beneficiary tailings are returned to slurry mixing and merged into the primary beneficiation operation;
(4)将步骤(3)所得的铅锌浮选混合精矿调浆至矿浆的液固质量比为3:1,在温度为60℃、硫酸浓度为95g/L的搅拌条件下,进行硫酸浸出90min,固液分离后得到含铅浸出渣和含锌浸出液;(4) the lead-zinc flotation mixed concentrate obtained in step (3) is slurried to a liquid-solid mass ratio of 3:1, and the sulfuric acid is carried out under the stirring condition that the temperature is 60° C. and the sulfuric acid concentration is 95 g/L. After leaching for 90 minutes, lead-containing leaching residue and zinc-containing leaching solution are obtained after solid-liquid separation;
(5)将步骤(4)所得的含铅浸出渣调浆至矿浆的液固质量比为3:1,在温度为50℃、甲基磺酸浓度为75g/L的搅拌条件下,进行甲基磺酸浸出45min,固液分离后得到含铅溶液和浸出尾渣;(5) the liquid-solid mass ratio of the lead-containing leaching slag obtained in step (4) to slurry is 3:1, and the temperature is 50 ° C and the methanesulfonic acid concentration is under the stirring condition of 75 g/L. The base sulfonic acid was leached for 45 minutes, and the lead-containing solution and the leaching tailings were obtained after solid-liquid separation;
本实施例中铅的回收率为88.2%,锌的回收率为89.1%。In this example, the recovery rate of lead was 88.2%, and the recovery rate of zinc was 89.1%.
实施例3:本实施例的微细粒嵌布型氧化铅锌矿中铅的质量百分数含量为6.7%,锌的质量百分数含量为7.3%;Embodiment 3: the mass percentage content of lead in the fine-grained embedded lead-zinc oxide ore of this embodiment is 6.7%, and the mass percentage content of zinc is 7.3%;
如图1所示,一种微细粒嵌布型氧化铅锌矿选冶分离与利用的方法,具体步骤如下:As shown in Figure 1, a kind of method for beneficiation and smelting separation and utilization of fine-grain embedded type lead-zinc oxide ore, the concrete steps are as follows:
(1)将微细粒嵌布型氧化铅锌矿破碎、磨细至-74μm粒级的质量百分含量占80%,调浆至矿浆质量百分浓度为40%,以每吨微细粒嵌布型氧化铅锌矿计,在矿浆中依次加入2200g组合抑制剂、900g铜氨络合物A、7600g组合硫化剂、600g铜氨络合物B、1400g组合捕收剂和80g起泡剂,进行粗选作业得到粗选精矿和粗选尾矿;(1) The fine-grained embedded lead-zinc oxide ore is crushed and ground to -74 μm with a mass percentage of 80%, and the pulp is adjusted to a mass percentage concentration of 40%. Type lead-zinc oxide ore, add 2200g combined inhibitor, 900g cupro-ammine complex A, 7600g combined vulcanizing agent, 600g cupro-ammine complex B, 1400g combined collector and 80g foaming agent in the pulp in sequence, and carry out The roughing operation produces rougher concentrate and rougher tailings;
(2)以每吨微细粒嵌布型氧化铅锌矿计,在步骤(1)粗选尾矿中依次加入1100g组合抑制剂、450g铜氨络合物A、3800g组合硫化剂、300g铜氨络合物B、700g组合捕收剂和40g起泡剂,进行一次扫选作业得到一次扫选精矿和一次扫选尾矿,其中一次扫选精矿返回调浆并入粗选作业;在一次扫选尾矿中依次加入550g组合抑制剂、225g铜氨络合物A、1900g组合硫化剂、150g铜氨络合物B、350g组合捕收剂和20g起泡剂,进行二次扫选作业得到二次扫选精矿和浮选尾矿,其中二次扫选精矿返回调浆并入一次扫选作业;(2) In terms of per ton of fine-grained embedded lead-zinc oxide ore, in step (1) roughing tailings were sequentially added 1100g combined inhibitor, 450g copper ammonia complex A, 3800g combined vulcanizing agent, 300g copper ammonia Complex B, 700g of combined collector and 40g of foaming agent, carry out a sweeping operation to obtain a sweeping concentrate and a sweeping tailings, of which the first sweeping concentrate is returned to slurry mixing and is incorporated into the roughing operation; 550g combined inhibitor, 225g copper ammonia complex A, 1900g combined vulcanizing agent, 150g copper ammonia complex B, 350g combined collector and 20g foaming agent were sequentially added to the tailings in the first sweep for secondary sweep. The operation obtains secondary scavenging concentrates and flotation tailings, of which the secondary scavenging concentrates are returned to slurry and merged into the primary scavenging operation;
(3)将步骤(1)粗选精矿进行一次精选作业得到一次精选精矿和一次精选尾矿,一次精选尾矿返回调浆并入粗选作业,一次精选精矿进行二次精选作业得到铅锌浮选混合精矿和二次精选尾矿,其中二次精选尾矿返回调浆并入一次精选作业;(3) Perform a beneficiation operation on the rougher concentrate in step (1) to obtain a beneficiary concentrate and a beneficiary tailings, a beneficiation tailings are returned to the slurry to be incorporated into the roughing operation, and a beneficiation concentrate is carried out The secondary beneficiation operation obtains lead-zinc flotation mixed concentrate and secondary beneficiation tailings, of which the secondary beneficiary tailings are returned to slurry mixing and merged into the primary beneficiation operation;
(4)将步骤(3)所得的铅锌浮选混合精矿调浆至矿浆的液固质量比为3:1,在温度为75℃、硫酸浓度为135g/L的搅拌条件下,进行硫酸浸出60min,固液分离后得到含铅浸出渣和含锌浸出液;(4) the lead-zinc flotation mixed concentrate obtained in step (3) is slurried to a liquid-solid mass ratio of 3:1, and the sulfuric acid is carried out under the stirring condition that the temperature is 75° C. and the sulfuric acid concentration is 135 g/L. After leaching for 60min, lead-containing leaching residue and zinc-containing leaching solution are obtained after solid-liquid separation;
(5)将步骤(4)所得的含铅浸出渣调浆至矿浆的液固质量比为3:1,在温度为60℃、甲基磺酸浓度为95g/L的搅拌条件下,进行甲基磺酸浸出35min,固液分离后得到含铅溶液和浸出尾渣;(5) the liquid-solid mass ratio of the lead-containing leaching slag of step (4) gained to slurry is 3:1, under the stirring condition that temperature is 60 DEG C, the methanesulfonic acid concentration is 95g/L, carry out methyl acid The base sulfonic acid was leached for 35 minutes, and the lead-containing solution and the leaching tailings were obtained after solid-liquid separation;
本实施例中铅的回收率为89.4%,锌的回收率为90.3%。In this example, the recovery rate of lead was 89.4%, and the recovery rate of zinc was 90.3%.
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Cited By (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN114908246A (en) * | 2021-02-10 | 2022-08-16 | 郑州大学 | A method for comprehensive dressing and smelting oxidation and sulfide lead-zinc mixed ore |
| CN117123372A (en) * | 2023-08-28 | 2023-11-28 | 中南大学 | Lead zinc oxide ore flotation combined activator and application thereof |
Citations (8)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101245407A (en) * | 2008-03-19 | 2008-08-20 | 云南冶金集团总公司 | Selecting-smelting method for processing complex hard-washing low ore grade pulmbous sulfide zincium mine |
| CN101816979A (en) * | 2010-03-16 | 2010-09-01 | 昆明理工大学 | Flotation activating agent of marmatite and blende and preparation method thereof |
| CN102002602A (en) * | 2010-11-10 | 2011-04-06 | 云南永昌铅锌股份有限公司 | Method for increasing recovery rates of lead and zinc during lead-zinc sulfide ore dressing-smelting joint application |
| 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 |
| CN105018726A (en) * | 2015-08-18 | 2015-11-04 | 云南驰宏锌锗股份有限公司 | Treatment method for lead and zinc paragenic ore |
| CN107185705A (en) * | 2017-05-23 | 2017-09-22 | 西北矿冶研究院 | Dressing and smelting combined recovery method for zinc in zinc oxide ore |
| CN108554618A (en) * | 2018-04-26 | 2018-09-21 | 昆明理工大学 | A kind of beneficiation method of copper/lead/zinc ore |
| CN111020176A (en) * | 2019-12-23 | 2020-04-17 | 昆明理工大学 | A method for combined recovery and utilization of copper-lead-zinc-oxygen-sulfur mixed ore dressing and smelting |
-
2020
- 2020-06-16 CN CN202010549651.1A patent/CN111659529B/en active Active
Patent Citations (8)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101245407A (en) * | 2008-03-19 | 2008-08-20 | 云南冶金集团总公司 | Selecting-smelting method for processing complex hard-washing low ore grade pulmbous sulfide zincium mine |
| CN101816979A (en) * | 2010-03-16 | 2010-09-01 | 昆明理工大学 | Flotation activating agent of marmatite and blende and preparation method thereof |
| CN102002602A (en) * | 2010-11-10 | 2011-04-06 | 云南永昌铅锌股份有限公司 | Method for increasing recovery rates of lead and zinc during lead-zinc sulfide ore dressing-smelting joint application |
| 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 |
| CN105018726A (en) * | 2015-08-18 | 2015-11-04 | 云南驰宏锌锗股份有限公司 | Treatment method for lead and zinc paragenic ore |
| CN107185705A (en) * | 2017-05-23 | 2017-09-22 | 西北矿冶研究院 | Dressing and smelting combined recovery method for zinc in zinc oxide ore |
| CN108554618A (en) * | 2018-04-26 | 2018-09-21 | 昆明理工大学 | A kind of beneficiation method of copper/lead/zinc ore |
| CN111020176A (en) * | 2019-12-23 | 2020-04-17 | 昆明理工大学 | A method for combined recovery and utilization of copper-lead-zinc-oxygen-sulfur mixed ore dressing and smelting |
Non-Patent Citations (2)
| Title |
|---|
| 中国有色金属学会: "《2016-2017 矿物加工工程 学科发展报告》", 31 March 2018 * |
| 丰奇成: "白铅矿氯离子强化硫化浮选试验及机理研究", 《中国博士学位论文全文数据库工程科技I辑》 * |
Cited By (3)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN114908246A (en) * | 2021-02-10 | 2022-08-16 | 郑州大学 | A method for comprehensive dressing and smelting oxidation and sulfide lead-zinc mixed ore |
| CN114908246B (en) * | 2021-02-10 | 2023-09-15 | 郑州大学 | A method for comprehensive selection and smelting of oxidation and lead-zinc sulfide mixed ores |
| CN117123372A (en) * | 2023-08-28 | 2023-11-28 | 中南大学 | Lead zinc oxide ore flotation combined activator and application thereof |
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