CN109351465B - Method for comprehensively recovering multiple metal elements in gold concentrate through step separation - Google Patents
Method for comprehensively recovering multiple metal elements in gold concentrate through step separation Download PDFInfo
- Publication number
- CN109351465B CN109351465B CN201811220161.6A CN201811220161A CN109351465B CN 109351465 B CN109351465 B CN 109351465B CN 201811220161 A CN201811220161 A CN 201811220161A CN 109351465 B CN109351465 B CN 109351465B
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- China
- Prior art keywords
- gold
- concentrate
- sulfur
- zinc
- ore pulp
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- 229910052737 gold Inorganic materials 0.000 title claims abstract description 239
- 239000010931 gold Substances 0.000 title claims abstract description 239
- PCHJSUWPFVWCPO-UHFFFAOYSA-N gold Chemical compound [Au] PCHJSUWPFVWCPO-UHFFFAOYSA-N 0.000 title claims abstract description 204
- 239000012141 concentrate Substances 0.000 title claims abstract description 153
- 238000000034 method Methods 0.000 title claims abstract description 95
- 238000000926 separation method Methods 0.000 title claims abstract description 41
- 229910052751 metal Inorganic materials 0.000 title claims abstract description 19
- 229910052717 sulfur Inorganic materials 0.000 claims abstract description 126
- 239000011593 sulfur Substances 0.000 claims abstract description 126
- 238000002386 leaching Methods 0.000 claims abstract description 121
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 claims abstract description 64
- 229910052709 silver Inorganic materials 0.000 claims abstract description 64
- 239000004332 silver Substances 0.000 claims abstract description 64
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims abstract description 56
- 229910052802 copper Inorganic materials 0.000 claims abstract description 56
- 239000010949 copper Substances 0.000 claims abstract description 56
- 238000005188 flotation Methods 0.000 claims abstract description 51
- JQJCSZOEVBFDKO-UHFFFAOYSA-N lead zinc Chemical compound [Zn].[Pb] JQJCSZOEVBFDKO-UHFFFAOYSA-N 0.000 claims abstract description 46
- 230000008569 process Effects 0.000 claims abstract description 39
- 238000003825 pressing Methods 0.000 claims abstract description 29
- 239000002253 acid Substances 0.000 claims abstract description 24
- 238000000605 extraction Methods 0.000 claims abstract description 15
- 239000012190 activator Substances 0.000 claims abstract description 4
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims description 130
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 claims description 79
- 239000003795 chemical substances by application Substances 0.000 claims description 67
- 229910052742 iron Inorganic materials 0.000 claims description 67
- 238000002156 mixing Methods 0.000 claims description 53
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 claims description 48
- OKTJSMMVPCPJKN-UHFFFAOYSA-N Carbon Chemical compound [C] OKTJSMMVPCPJKN-UHFFFAOYSA-N 0.000 claims description 40
- 229910052799 carbon Inorganic materials 0.000 claims description 40
- 229910052725 zinc Inorganic materials 0.000 claims description 36
- 239000011701 zinc Substances 0.000 claims description 36
- 239000007788 liquid Substances 0.000 claims description 35
- 239000012065 filter cake Substances 0.000 claims description 28
- 239000004088 foaming agent Substances 0.000 claims description 27
- 230000002000 scavenging effect Effects 0.000 claims description 24
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims description 21
- 239000000706 filtrate Substances 0.000 claims description 21
- 229910052760 oxygen Inorganic materials 0.000 claims description 21
- 239000001301 oxygen Substances 0.000 claims description 21
- 238000003756 stirring Methods 0.000 claims description 19
- 239000003814 drug Substances 0.000 claims description 16
- 230000003213 activating effect Effects 0.000 claims description 15
- 238000006073 displacement reaction Methods 0.000 claims description 14
- 239000007787 solid Substances 0.000 claims description 14
- 238000004064 recycling Methods 0.000 claims description 12
- 238000005406 washing Methods 0.000 claims description 12
- 229910000365 copper sulfate Inorganic materials 0.000 claims description 11
- ARUVKPQLZAKDPS-UHFFFAOYSA-L copper(II) sulfate Chemical compound [Cu+2].[O-][S+2]([O-])([O-])[O-] ARUVKPQLZAKDPS-UHFFFAOYSA-L 0.000 claims description 11
- TUZCOAQWCRRVIP-UHFFFAOYSA-N butoxymethanedithioic acid Chemical compound CCCCOC(S)=S TUZCOAQWCRRVIP-UHFFFAOYSA-N 0.000 claims description 10
- 239000000203 mixture Substances 0.000 claims description 10
- 238000003723 Smelting Methods 0.000 claims description 9
- 239000002994 raw material Substances 0.000 claims description 9
- 238000001179 sorption measurement Methods 0.000 claims description 9
- 239000003513 alkali Substances 0.000 claims description 8
- 238000000746 purification Methods 0.000 claims description 8
- 229910000831 Steel Inorganic materials 0.000 claims description 7
- 238000006243 chemical reaction Methods 0.000 claims description 7
- 238000003795 desorption Methods 0.000 claims description 7
- 238000005868 electrolysis reaction Methods 0.000 claims description 7
- 238000000227 grinding Methods 0.000 claims description 7
- 230000001590 oxidative effect Effects 0.000 claims description 7
- 230000009467 reduction Effects 0.000 claims description 7
- 239000010959 steel Substances 0.000 claims description 7
- HQABUPZFAYXKJW-UHFFFAOYSA-O butylazanium Chemical compound CCCC[NH3+] HQABUPZFAYXKJW-UHFFFAOYSA-O 0.000 claims description 6
- 238000011085 pressure filtration Methods 0.000 claims description 6
- GCLGEJMYGQKIIW-UHFFFAOYSA-H sodium hexametaphosphate Chemical compound [Na]OP1(=O)OP(=O)(O[Na])OP(=O)(O[Na])OP(=O)(O[Na])OP(=O)(O[Na])OP(=O)(O[Na])O1 GCLGEJMYGQKIIW-UHFFFAOYSA-H 0.000 claims description 6
- 235000019982 sodium hexametaphosphate Nutrition 0.000 claims description 6
- 239000001577 tetrasodium phosphonato phosphate Substances 0.000 claims description 6
- XFXPMWWXUTWYJX-UHFFFAOYSA-N Cyanide Chemical compound N#[C-] XFXPMWWXUTWYJX-UHFFFAOYSA-N 0.000 claims description 5
- XEIPQVVAVOUIOP-UHFFFAOYSA-N [Au]=S Chemical compound [Au]=S XEIPQVVAVOUIOP-UHFFFAOYSA-N 0.000 claims description 5
- 239000002002 slurry Substances 0.000 claims description 5
- HRZFUMHJMZEROT-UHFFFAOYSA-L sodium disulfite Chemical compound [Na+].[Na+].[O-]S(=O)S([O-])(=O)=O HRZFUMHJMZEROT-UHFFFAOYSA-L 0.000 claims description 5
- 235000010262 sodium metabisulphite Nutrition 0.000 claims description 5
- 239000012991 xanthate Substances 0.000 claims description 5
- 241000544912 Melanoides Species 0.000 claims description 4
- 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 description 3
- 238000007654 immersion Methods 0.000 claims description 3
- 239000011734 sodium Substances 0.000 claims description 3
- 229910052708 sodium Inorganic materials 0.000 claims description 3
- LFQSCWFLJHTTHZ-UHFFFAOYSA-N Ethanol Chemical compound CCO LFQSCWFLJHTTHZ-UHFFFAOYSA-N 0.000 claims description 2
- QTANTQQOYSUMLC-UHFFFAOYSA-O Ethidium cation Chemical compound C12=CC(N)=CC=C2C2=CC=C(N)C=C2[N+](CC)=C1C1=CC=CC=C1 QTANTQQOYSUMLC-UHFFFAOYSA-O 0.000 claims description 2
- 229910002651 NO3 Inorganic materials 0.000 claims description 2
- NHNBFGGVMKEFGY-UHFFFAOYSA-N Nitrate Chemical compound [O-][N+]([O-])=O NHNBFGGVMKEFGY-UHFFFAOYSA-N 0.000 claims description 2
- 235000008331 Pinus X rigitaeda Nutrition 0.000 claims description 2
- 235000011613 Pinus brutia Nutrition 0.000 claims description 2
- 241000018646 Pinus brutia Species 0.000 claims description 2
- 230000035484 reaction time Effects 0.000 claims description 2
- 238000005516 engineering process Methods 0.000 abstract description 11
- 238000011084 recovery Methods 0.000 abstract description 8
- 230000008901 benefit Effects 0.000 abstract description 7
- 239000003153 chemical reaction reagent Substances 0.000 abstract description 5
- 230000007613 environmental effect Effects 0.000 abstract description 3
- 235000020893 clean eating Nutrition 0.000 abstract description 2
- 238000001035 drying Methods 0.000 abstract description 2
- 230000004913 activation Effects 0.000 abstract 1
- 238000002360 preparation method Methods 0.000 abstract 1
- 239000011133 lead Substances 0.000 description 26
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 description 25
- 239000003921 oil Substances 0.000 description 19
- CDBYLPFSWZWCQE-UHFFFAOYSA-L Sodium Carbonate Chemical compound [Na+].[Na+].[O-]C([O-])=O CDBYLPFSWZWCQE-UHFFFAOYSA-L 0.000 description 12
- KXZJHVJKXJLBKO-UHFFFAOYSA-N chembl1408157 Chemical compound N=1C2=CC=CC=C2C(C(=O)O)=CC=1C1=CC=C(O)C=C1 KXZJHVJKXJLBKO-UHFFFAOYSA-N 0.000 description 11
- 239000000047 product Substances 0.000 description 9
- VHUUQVKOLVNVRT-UHFFFAOYSA-N Ammonium hydroxide Chemical compound [NH4+].[OH-] VHUUQVKOLVNVRT-UHFFFAOYSA-N 0.000 description 6
- IJGRMHOSHXDMSA-UHFFFAOYSA-N Atomic nitrogen Chemical compound N#N IJGRMHOSHXDMSA-UHFFFAOYSA-N 0.000 description 6
- 235000008733 Citrus aurantifolia Nutrition 0.000 description 6
- 235000011941 Tilia x europaea Nutrition 0.000 description 6
- 235000011114 ammonium hydroxide Nutrition 0.000 description 6
- 239000003818 cinder Substances 0.000 description 6
- 239000004571 lime Substances 0.000 description 6
- 239000002184 metal Substances 0.000 description 6
- 150000002739 metals Chemical class 0.000 description 6
- 235000017550 sodium carbonate Nutrition 0.000 description 6
- 229910000029 sodium carbonate Inorganic materials 0.000 description 6
- UGFAIRIUMAVXCW-UHFFFAOYSA-N Carbon monoxide Chemical compound [O+]#[C-] UGFAIRIUMAVXCW-UHFFFAOYSA-N 0.000 description 5
- 239000003546 flue gas Substances 0.000 description 5
- 239000012535 impurity Substances 0.000 description 4
- 229910052500 inorganic mineral Inorganic materials 0.000 description 4
- 238000004519 manufacturing process Methods 0.000 description 4
- 239000011707 mineral Substances 0.000 description 4
- 235000010755 mineral Nutrition 0.000 description 4
- 239000002699 waste material Substances 0.000 description 4
- 241000196324 Embryophyta Species 0.000 description 3
- 239000007789 gas Substances 0.000 description 3
- 229910052757 nitrogen Inorganic materials 0.000 description 3
- 230000003647 oxidation Effects 0.000 description 3
- 238000007254 oxidation reaction Methods 0.000 description 3
- 239000000843 powder Substances 0.000 description 3
- CSCPPACGZOOCGX-UHFFFAOYSA-N Acetone Chemical compound CC(C)=O CSCPPACGZOOCGX-UHFFFAOYSA-N 0.000 description 2
- WUOACPNHFRMFPN-UHFFFAOYSA-N alpha-terpineol Chemical compound CC1=CCC(C(C)(C)O)CC1 WUOACPNHFRMFPN-UHFFFAOYSA-N 0.000 description 2
- 230000007797 corrosion Effects 0.000 description 2
- 238000005260 corrosion Methods 0.000 description 2
- 238000005536 corrosion prevention Methods 0.000 description 2
- SQIFACVGCPWBQZ-UHFFFAOYSA-N delta-terpineol Natural products CC(C)(O)C1CCC(=C)CC1 SQIFACVGCPWBQZ-UHFFFAOYSA-N 0.000 description 2
- 238000009826 distribution Methods 0.000 description 2
- 239000006260 foam Substances 0.000 description 2
- LELOWRISYMNNSU-UHFFFAOYSA-N hydrogen cyanide Chemical compound N#C LELOWRISYMNNSU-UHFFFAOYSA-N 0.000 description 2
- 239000003112 inhibitor Substances 0.000 description 2
- 239000010665 pine oil Substances 0.000 description 2
- 229910052683 pyrite Inorganic materials 0.000 description 2
- NIFIFKQPDTWWGU-UHFFFAOYSA-N pyrite Chemical compound [Fe+2].[S-][S-] NIFIFKQPDTWWGU-UHFFFAOYSA-N 0.000 description 2
- 239000011028 pyrite Substances 0.000 description 2
- 229940001584 sodium metabisulfite Drugs 0.000 description 2
- 229910052569 sulfide mineral Inorganic materials 0.000 description 2
- 150000004763 sulfides Chemical class 0.000 description 2
- 229940116411 terpineol Drugs 0.000 description 2
- PAWQVTBBRAZDMG-UHFFFAOYSA-N 2-(3-bromo-2-fluorophenyl)acetic acid Chemical compound OC(=O)CC1=CC=CC(Br)=C1F PAWQVTBBRAZDMG-UHFFFAOYSA-N 0.000 description 1
- BITVSRNAFFZUFW-UHFFFAOYSA-N 5-ethyl-6-phenylphenanthridin-5-ium-3,8-diamine;chloride Chemical compound [Cl-].C12=CC(N)=CC=C2C2=CC=C(N)C=C2[N+](CC)=C1C1=CC=CC=C1 BITVSRNAFFZUFW-UHFFFAOYSA-N 0.000 description 1
- IRDQNLLVRXMERV-UHFFFAOYSA-N CCCC[Na] Chemical compound CCCC[Na] IRDQNLLVRXMERV-UHFFFAOYSA-N 0.000 description 1
- ZKQDCIXGCQPQNV-UHFFFAOYSA-N Calcium hypochlorite Chemical compound [Ca+2].Cl[O-].Cl[O-] ZKQDCIXGCQPQNV-UHFFFAOYSA-N 0.000 description 1
- PFRUBEOIWWEFOL-UHFFFAOYSA-N [N].[S] Chemical compound [N].[S] PFRUBEOIWWEFOL-UHFFFAOYSA-N 0.000 description 1
- 238000004458 analytical method Methods 0.000 description 1
- 230000001580 bacterial effect Effects 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 239000007844 bleaching agent Substances 0.000 description 1
- 125000004122 cyclic group Chemical group 0.000 description 1
- 238000005262 decarbonization Methods 0.000 description 1
- 229940079593 drug Drugs 0.000 description 1
- 238000005265 energy consumption Methods 0.000 description 1
- 238000003912 environmental pollution Methods 0.000 description 1
- ZSFDBVJMDCMTBM-UHFFFAOYSA-N ethane-1,2-diamine;phosphoric acid Chemical compound NCCN.OP(O)(O)=O ZSFDBVJMDCMTBM-UHFFFAOYSA-N 0.000 description 1
- ZOOODBUHSVUZEM-UHFFFAOYSA-N ethoxymethanedithioic acid Chemical compound CCOC(S)=S ZOOODBUHSVUZEM-UHFFFAOYSA-N 0.000 description 1
- 239000011790 ferrous sulphate Substances 0.000 description 1
- 235000003891 ferrous sulphate Nutrition 0.000 description 1
- 238000001914 filtration Methods 0.000 description 1
- 239000008396 flotation agent Substances 0.000 description 1
- 238000010438 heat treatment Methods 0.000 description 1
- 229910001385 heavy metal Inorganic materials 0.000 description 1
- 238000007731 hot pressing Methods 0.000 description 1
- 230000002401 inhibitory effect Effects 0.000 description 1
- 150000002500 ions Chemical class 0.000 description 1
- OVMJVEMNBCGDGM-UHFFFAOYSA-N iron silver Chemical compound [Fe].[Ag] OVMJVEMNBCGDGM-UHFFFAOYSA-N 0.000 description 1
- BAUYGSIQEAFULO-UHFFFAOYSA-L iron(2+) sulfate (anhydrous) Chemical compound [Fe+2].[O-]S([O-])(=O)=O BAUYGSIQEAFULO-UHFFFAOYSA-L 0.000 description 1
- 229910000359 iron(II) sulfate Inorganic materials 0.000 description 1
- 239000003350 kerosene Substances 0.000 description 1
- 238000011031 large-scale manufacturing process Methods 0.000 description 1
- 230000007774 longterm Effects 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 150000002825 nitriles Chemical class 0.000 description 1
- 235000019353 potassium silicate Nutrition 0.000 description 1
- 239000010970 precious metal Substances 0.000 description 1
- NTHWMYGWWRZVTN-UHFFFAOYSA-N sodium silicate Chemical compound [Na+].[Na+].[O-][Si]([O-])=O NTHWMYGWWRZVTN-UHFFFAOYSA-N 0.000 description 1
- AAJRIJBGDLLRAE-UHFFFAOYSA-M sodium;butoxymethanedithioate Chemical compound [Na+].CCCCOC([S-])=S AAJRIJBGDLLRAE-UHFFFAOYSA-M 0.000 description 1
- 239000000126 substance Substances 0.000 description 1
- 239000001117 sulphuric acid Substances 0.000 description 1
- 235000011149 sulphuric acid Nutrition 0.000 description 1
- 239000001648 tannin Substances 0.000 description 1
- 235000018553 tannin Nutrition 0.000 description 1
- 229920001864 tannin Polymers 0.000 description 1
- YBRBMKDOPFTVDT-UHFFFAOYSA-N tert-butylamine Chemical compound CC(C)(C)N YBRBMKDOPFTVDT-UHFFFAOYSA-N 0.000 description 1
Classifications
-
- 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
- B03B7/00—Combinations of wet processes or apparatus with other processes or apparatus, e.g. for dressing ores or garbage
-
- 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
-
- 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
- B03D1/004—Organic compounds
- B03D1/014—Organic compounds containing phosphorus
-
- 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
- B03D1/004—Organic compounds
- B03D1/016—Macromolecular compounds
-
- 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
-
- 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
- C22B11/08—Obtaining noble metals by cyaniding
-
- 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
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/12—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic alkaline solutions
-
- 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
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
- C22B3/44—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes
- C22B3/46—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes by substitution, e.g. by cementation
-
- 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
-
- 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
- B03D2201/04—Frothers
-
- 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
- B03D2203/025—Precious metal ores
-
- 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)
- Engineering & Computer Science (AREA)
- Manufacturing & Machinery (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Metallurgy (AREA)
- Organic Chemistry (AREA)
- Life Sciences & Earth Sciences (AREA)
- Inorganic Chemistry (AREA)
- Environmental & Geological Engineering (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Geochemistry & Mineralogy (AREA)
- Geology (AREA)
- Chemical Kinetics & Catalysis (AREA)
- General Chemical & Material Sciences (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention relates to a method for comprehensively recovering multiple metal elements in gold concentrate by gradient separation, wherein the first-stage process comprises the step of carrying out pre-flotation technology on low-sulfur gold concentrate by using a high-efficiency mixed activator and a flotation reagent, the second-stage process comprises the step of carrying out alkaline leaching cyanidation gold extraction technology on the high-sulfur concentrate, the third-stage process comprises the step of carrying out environment-friendly reagent wet leaching gold and silver technology on the low-sulfur tailing, the fourth-stage process comprises the step of carrying out high-sulfur gold concentrate cyanidation tailing, namely the first low-gold high-sulfur concentrate preferential lead-zinc ore dressing technology, the fifth-stage process comprises the step of carrying out lead-zinc tailing activation copper dressing technology, the sixth-stage process comprises the step of roasting copper dressing tailing, namely the step of carrying out acid preparation and enrichment gold and gold; the process prolongs the chain for comprehensively utilizing the low-sulfur gold concentrate resources, improves the comprehensive recovery value of the resources and the grade quality of the products, realizes the clean eating, drying and pressing of the gold concentrate resources, and has great economic benefit, environmental protection benefit and social benefit.
Description
Technical Field
The invention relates to a method for recovering metal elements from gold concentrate, in particular to a method for recovering the metal elements from the gold concentrate through gradient separation, and belongs to the technical field of flotation, gold smelting and comprehensive recovery and utilization of mineral resources.
Background
The gold smelting industry generally adopts a cyaniding gold extraction method which is divided into a direct cyaniding method, a roasting cyaniding method, a two-section roasting cyaniding method, a bacterial oxidation cyaniding method and a hot-pressing oxidation cyaniding method, wherein except the direct cyaniding method, the method is a pretreatment process for eliminating the harm to gold extraction, the influence of harmful impurities is effectively eliminated through the pretreatment process, and the leaching rate of gold and silver is improved. The main gold extraction method is a direct cyanidation method, gold concentrate which has few harmful elements and high gold and silver leaching rate and is suitable for direct cyanidation treatment is subjected to direct cyanidation to extract gold and silver, generated cyanidation tailings are stockpiled or used as raw materials of a sulfuric acid plant for roasting to prepare acid, although sulfur in the tailings is utilized, valuable metals such as lead, copper and iron contained in the tailings cannot be effectively recovered, and in addition, because the cyanidation tailings contain some heavy metal ions, cyanides and residual flotation reagents, the long-term mass stockpiling can cause environmental pollution.
In recent years, many researchers have studied on the recovery of valuable metals from cyanidation tailings, but the cyanidation tailings are added with agents such as sodium cyanide, lime, bleaching powder and the like in the production process, so that after long-time cyanidation treatment, the properties of the surface of minerals are obviously changed, in particular, polymetallic pyrite and other sulfides are easily oxidized to reduce the floatability of the polymetallic pyrite and other sulfides, and the cyanide strongly inhibits the sulfide minerals, so that the cyanidation tailings which are difficult to beneficiate per se are more difficult to float. Because the cyaniding tailings often contain more residual reagents, the interference on flotation is large, the properties of ores are relatively complex, the granularity of the ores is extremely fine, and the finding of a proper flotation reagent is the biggest problem at present. Researchers have made much research on agents for flotation and gold extraction of cyanidation tailings, collecting agents are mainly concentrated on xanthate, black and sulfur nitrogen collecting agents, and inhibitors are mainly concentrated on water glass, sodium hexametaphosphate, tannin and the like. The existing research and analysis finds that the conventional single collecting agent is difficult to simultaneously consider selectivity and collecting capacity, and the problem of non-ideal inhibiting effect exists in the case of single inhibitor.
The invention patent with patent number 201410476017.4 discloses a method for removing harmful impurities before cyanidation tailings are used for flotation of high-grade sulfur concentrate, and the invention patent discloses: the acid leaching tanks and the stirring devices in the tanks are connected in series, so that the acid leaching time is ensured to be 100-150 min. The ore pulp passes through each acid leaching groove in proper order before the flotation to add dilute sulphuric acid in first acid leaching groove, all the other three grooves add the edulcoration medicament, and each acid leaching groove lets in steam respectively, raises the temperature for harmful impurity volatilizees, and the gas that volatilizees gets into tail gas absorbing device, and the ore pulp enters into the flotation machine after getting rid of harmful impurity. The invention highlights the earlier stage treatment method of high-grade sulfur concentrate flotation from cyanidation tailings, the flow is extremely complex, the steam heating method is adopted, the method has practicability for a small-sized production system, and the method has high energy consumption, high cost and no operability for processing a large-scale production processing factory.
The invention patent No. 201410678782.4 discloses a method for recovering gold from cyanided tailings by branch series flow flotation, which mainly comprises the following steps: pretreating ore pulp, namely adopting sodium carbonate and ferrous sulfate as an ore pulp regulator; carrying out flotation and carbon removal on kerosene; shunting and mixing slurry: the ore pulp after carbon removal automatically flows to a stirring distribution tank, and the ore pulp is averagely distributed to two pulp mixing barrels by the stirring distribution tank for pulp mixing; branch series flow flotation: the foam products of the first roughing step are all subjected to flotation together with the ore pulp after the decarbonization, the foam products of the second roughing step are subjected to closed-circuit concentration operation of a first concentration step and a second concentration step to obtain gold concentrate, and no medicament is added in the concentration step; and (4) performing closed scavenging on the product in the first roughing tank by a first scavenging and a second scavenging to obtain tailings. The applicability of the invention is only limited to full-mud cyanidation tailings, and the carbon-containing gold concentrate and the gold concentrate are produced by adopting a step-by-step flotation process, the flow is complex, and the further treatment and purification method for treating the gold concentrate produced by enrichment and the carbon-containing gold concentrate cannot be explained in detail.
The invention discloses a method for recovering gold by flotation of cyaniding gold leaching tailings, which is disclosed by the patent No. 201710913404.3, and mainly aims at the current situation that the cyaniding gold leaching residues are subjected to pre-oxidation treatment on raw ores before cyaniding leaching, the surface properties of gold carrier minerals are greatly changed, and enrichment recovery cannot be performed on the gold carrier minerals by a conventional flotation method. The applicability of the invention is only limited to cyanide tailings which are subjected to preoxidation treatment before cyaniding leaching, a plurality of agents such as sulfuric acid, ethylenediamine phosphate, acetone and the like are added in the process to float out organic carbon and sulfide minerals carrying gold respectively, the process flow is complex, the agents are added in various types, particularly, the sulfuric acid is added, cyanide is decomposed to generate volatile hydrogen cyanide gas which causes harm to the environment, in addition, the sulfuric acid is added to cause corrosion to equipment, the corrosion prevention cost is increased due to the enhancement of the corrosion prevention of the equipment, and the service life of the equipment is still influenced.
The comprehensive recycling of resources is surrounded, the high-gold high-sulfur ore and the low-gold low-sulfur tailing are produced by adopting a pre-flotation process to the sulfur-containing gold concentrate, the direct cyanidation gold extraction process is adopted to the high-gold high-sulfur ore, the gold and silver precious metals are comprehensively recycled, the lead-zinc bulk concentrate and the copper concentrate are comprehensively recycled by adopting a secondary flotation process to the high-sulfur ore after direct cyanidation, the lead-zinc bulk concentrate can be directly sold to the outside, the comprehensive utilization of resources is realized, the secondary flotation tailing is gold-containing sulfur concentrate containing more than 48% of gold 1-1.5g/t, the gold-containing sulfur concentrate is roasted to prepare acid, the sulfuric acid cinder is high-iron cinder, the total iron content is more than 62%, and the total iron content can be directly sold to the outside. The grade of gold in the high-iron cinder generated by roasting reaches more than 2g/t, and the grade of silver in the high-iron cinder reaches more than 10g/t, so that the high-iron cinder has high recovery value. Therefore, the method preferentially recovers the gold and silver in the generated high-iron cinder, discusses the comprehensive recovery of valuable metals to the maximum extent, and then sells the valuable metals as fine iron powder, generates greater economic benefit, and becomes an important subject of the industry.
Disclosure of Invention
The invention provides a method for comprehensively recovering multiple metal elements in gold concentrate by step separation, aiming at the current situation that valuable metals in existing cyanided tailings cannot be fully utilized.
The technical scheme for solving the technical problems is as follows:
a method for comprehensively recovering multiple metal elements in gold concentrate through step separation comprises the following steps:
1) pre-flotation: taking gold-containing 10-50g/t, silver-containing 10-80g/t, copper-containing 0.1-0.5 wt%, lead-containing 0.1-0.5 wt%, zinc-containing 0.1-0.5 wt%, sulfur-containing 20-40 wt% and low-sulfur gold concentrate with fineness of-400 meshes lower than 90% as raw materials, carrying out liquid mixing with a flotation system, grinding ore pulp until the fineness of-400 meshes reaches more than 90%, adding 1000-3000g/t of activating agent, 800-1200g/t of flotation collecting agent and 100-200g/t of foaming agent into the ore pulp, wherein the activating agent is one or a mixture of sodium pyrosulfite, sodium hexametaphosphate and copper sulfate, the flotation collecting agent is one or a mixture of sodium xanthates, 25# melanoides and butylammonium melanoides, and the foaming agent is one or a mixture of pine alcohol oil and 2# oil, through the processes of primary roughing, secondary scavenging and secondary concentration flotation, high gold-sulfur concentrate with gold grade of 20-80g/t, silver grade of 20-100g/t, copper grade of 0.2-0.8 wt%, lead grade of 0.2-0.8 wt%, zinc grade of 0.2-0.8 wt% and sulfur grade of not less than 48 wt% and low-sulfur tailings with gold grade of 1.0-6.0g/t, silver grade of 3-15g/t and sulfur grade of 1-6 wt% are produced;
2) alkaline leaching, cyaniding and gold extraction: the high gold and sulfur concentrate obtained in the step 1) is subjected to size mixing by adopting alkali immersion liquid, the pH value of ore pulp is adjusted to 9-12, and 30-50m is filled into the ore pulp3Air per hour is subjected to alkaline leaching reaction under the stirring condition, then alkaline leaching solution and alkaline leaching filter cake are obtained through pressure filtration, the obtained alkaline leaching solution is recycled, the obtained alkaline leaching filter cake is subjected to size mixing by using a first zinc powder replacement barren solution, then cyaniding leaching is carried out, the concentration of cyanide is controlled to be 0.10-0.60%, leached ore pulp is fully washed and subjected to liquid-solid separation to obtain gold-containing pregnant solution and a first low-gold high-sulfur concentrate, the gold-containing pregnant solution is subjected to purification, deoxidation, zinc powder replacement and plate-and-frame pressure filtration to obtain gold mud and first zinc powderDisplacing the barren solution;
3) leaching gold and silver from low-sulfur tailings by a wet method: mixing the low-sulfur tailings obtained in the step 1) with a second zinc powder displacement barren solution, adjusting the pH value of ore pulp to 8-10, adding 2-6kg/t of environment-friendly medicament HBJ serving as a leaching agent into the ore pulp, fully washing the leached ore pulp and performing liquid-solid separation to obtain gold-containing pregnant solution and low-sulfur tailings, and purifying, deoxidizing, zinc powder displacement and plate-and-frame filter pressing the gold-containing pregnant solution to obtain gold mud and a second zinc powder displacement barren solution;
4) lead and zinc ore selection: mixing the first low-gold high-sulfur concentrate generated in the step 2) by adopting a lead concentration system, then adjusting the pH value of the ore pulp to be 8-10, adding 300-600g/t of copper sulfate as a zinc activator, simultaneously adding 80-120g/t of lead-zinc ore dressing agent and 60-80g/t of foaming agent, wherein the lead-zinc ore dressing agent is one or a mixture of butyl xanthate and ethidium nitrate, the foaming agent is No. 2 oil, obtaining 20-30% of lead-zinc mixed concentrate and 20-30% of zinc by primary roughing and secondary scavenging, and selling the lead-zinc mixed concentrate to a lead-zinc smelting enterprise;
5) copper selection: carrying out size mixing on the lead-zinc tailings obtained in the step 4) by adopting a copper selecting system liquid, then adjusting the pH value of ore pulp to 5-7, adding 80-120g/t of copper selecting agent and 60-80g/t of foaming agent into the ore pulp, wherein the copper selecting agent is one or a mixture of butyl xanthate and butyl ammonium nigride, the foaming agent is 2# oil, and carrying out once roughing, twice scavenging and twice fine concentration operation to obtain copper concentrate containing 10-20 wt% of copper and second low-gold high-sulfur concentrate;
6) roasting to prepare acid and enrich gold, silver and iron to produce iron ore concentrate: feeding the second low-gold high-sulfur concentrate obtained in the step 5) into a fluidized bed roaster for roasting in a weak oxygen reduction atmosphere, wherein the oxygen supply amount is 85-90% of the oxygen demand of oxidizing roasting, the roasting temperature is controlled at 800 ℃ with 700-;
7) leaching gold and silver from iron ore concentrate by a wet method: the iron ore concentrate obtained in the step 6) is subjected to filter pressing by a filter press, filtrate is subjected to size mixing, the pH value of ore pulp is adjusted to 8-10, and the ore pulp is subjected to size mixingAdding 2-6kg/t environment-friendly agent HBJ as leaching agent, adsorbing the leached ore pulp by a carbon adsorption stirring tank with carbon density of 50-100kg/m3Ore pulp is screened by a carbon screen to obtain gold-loaded carbon and decarburized ore pulp, the gold-loaded carbon is subjected to desorption electrolysis process to obtain finished gold, the decarburized ore pulp is subjected to filter pressing by a filter press to obtain filter cakes and filtrate, the filtrate is returned to a system for recycling, and the filter cakes are iron ore concentrates with the sulfur content of less than or equal to 0.5 percent, the gold content of 0.5-0.8g/t, the silver content of 1-5g/t and the iron content of 62-65 percent and are sold to a steel mill;
wherein, the step 2) and the step 3) are not in sequence.
Further, the concentration of the ore pulp in the step 1) is controlled to be 50-70 wt%.
Further, in the step 2), the concentration of the ore pulp is controlled to be 20-50 wt% by alkali leaching solution size mixing, and the concentration of the ore pulp is controlled to be 10-50 wt% by using the first zinc powder to replace the barren solution size mixing.
Further, the stirring speed in the step 2) is 30-50r/min, and the alkaline leaching reaction time is 4-6 hours.
Further, in the step 3), the concentration of the ore pulp is controlled to be 10-50 wt%, and the leaching time is 8-24 hours.
Further, the concentration of the ore pulp is controlled to be 10-30 wt% in the step 4) and the step 5).
Further, in the step 7), the concentration of the ore pulp is controlled to be 30-50 wt%, the leaching time is 8-24 hours, and the time of the adsorption process is 8-16 hours.
The process has the beneficial effects that:
1) the first level of the process of the invention is a pre-flotation technology for low-sulfur gold concentrate by utilizing a high-efficiency mixed activating agent and a flotation agent, the second level is an alkaline leaching cyaniding gold extraction technology for high-sulfur concentrate, the third level is a technology for leaching gold and silver from low-sulfur tailings by adopting an environment-friendly agent wet method, the fourth level is a technology for preferentially separating lead and zinc from high-sulfur gold concentrate cyanided tailings, namely a first low-gold high-sulfur concentrate, the fifth level is a technology for activating and selecting copper from lead and zinc tailings, the sixth level is a technology for producing iron concentrate by roasting copper-selecting tailings, namely a second low-gold high-sulfur concentrate to produce acid and enrich gold and silver iron, and the seventh level is a technology for;
2) the process not only avoids and eliminates the technical problems of complex process flow, high production cost, short equipment corrosion service cycle, poor production continuity, unstable indexes and the like of valuable metals such as gold and silver and the like recovered from cyaniding gold extraction waste residues, but also comprehensively and integrally recovers the gold, the silver and the copper in low-sulfur gold concentrate by a step separation multistage technology, thereby achieving the purposes of comprehensive treatment and recovery, cyclic utilization, economy and environmental protection.
3) The process prolongs the chain for comprehensively utilizing the low-sulfur gold concentrate resources, improves the comprehensive recovery value of the resources and the grade quality of the products, realizes the clean eating, drying and pressing of the gold concentrate resources, and has great economic benefit, environmental protection benefit and social benefit.
Detailed Description
The principles and features of this invention are described below in conjunction with examples, which are set forth to illustrate, but are not to be construed to limit the scope of the invention.
Example 1:
a method for comprehensively recovering multiple metal elements in gold concentrate through step separation comprises the following steps:
1) pre-flotation: takes low-sulfur gold concentrate with 10g/t of gold, 10g/t of silver, 0.1% of copper, 0.1% of lead, 0.1% of zinc, 20% of sulfur and 68% of fineness of-400 meshes as raw materials, mixing the pulp with the liquid of a flotation system, controlling the concentration of the pulp to be 50 wt%, grinding the pulp to the fineness of-400 meshes and reaching more than 90%, adding 1000g/t of sodium metabisulfite activating agent, 800g/t of flotation collector sodium butyl xanthate and 100g/t of pine oil foaming agent, and performing primary roughing, secondary scavenging and secondary concentration flotation to produce high-gold sulfur concentrate with gold grade of 20g/t, silver grade of 20g/t, copper grade of 0.2 wt%, lead grade of 0.2 wt%, zinc grade of 0.2 wt% and sulfur grade of 48 wt% and low-sulfur tailings with gold grade of 1.0g/t, silver grade of 3g/t and sulfur grade of 1 wt%;
2) alkaline leaching, cyaniding and gold extraction: the high gold and sulfur concentrate obtained in the step 1) is subjected to size mixing by adopting alkali leaching solution until the concentration of ore pulp is 20 wt%, lime is added to adjust the pH value to be 9, and 30m is charged into the ore pulp3Air per hour, performing alkaline leaching reaction under stirring at 30r/min for 4 hr, and press filtering to obtain alkaline leaching solution and alkaline leachingThe method comprises the following steps of (1) recycling a filter cake, using an alkaline leaching filter cake to perform size mixing by using a first zinc powder displacement barren solution, controlling the concentration of ore pulp to be 10 wt%, performing cyanidation leaching, using sodium cyanide as a leaching agent, controlling the concentration of sodium cyanide to be 0.10%, wherein the leaching time is 36 hours, performing full washing and liquid-solid separation on the leached ore pulp to obtain gold-containing pregnant solution and a first low-gold high-sulfur concentrate, and performing purification, deoxidation, zinc powder displacement and plate and frame filter pressing on the gold-containing pregnant solution to obtain gold mud and a first zinc powder displacement barren solution;
3) leaching gold and silver from low-sulfur tailings by a wet method: mixing the low-sulfur tailings with gold grade of 1.0g/t, silver grade of 3g/t and sulfur grade of 1 wt% obtained in the step 1) with second zinc powder replacement barren solution, controlling the concentration of the ore pulp to be 10 wt%, adjusting the pH value of the ore pulp to be 8, adding 2kg/t of environment-friendly medicament HBJ serving as a leaching agent into the ore pulp, fully washing and carrying out liquid-solid separation on the leached ore pulp to obtain gold-containing pregnant solution and low-sulfur tailings, and purifying, deoxidizing, zinc powder replacement and plate-and-frame filter pressing the gold-containing pregnant solution to obtain gold mud and second zinc powder replacement barren solution;
4) lead and zinc ore selection: carrying out liquid size mixing on the first low-gold high-sulfur concentrate generated in the step 2) by using a lead-selecting system, then adjusting the pH value of the ore pulp to be 8 by using soda ash, adding 300g/t of copper sulfate as a zinc activating agent, simultaneously adding 80g/t of butyl xanthate as a lead-zinc ore dressing agent and 60g/t of No. 2 oil as a foaming agent, carrying out once roughing, twice scavenging and twice selecting operation to obtain lead-zinc mixed concentrate containing 20% of lead and 20% of zinc and lead-zinc tailings, and selling the lead-zinc mixed concentrate to a lead-zinc smelting enterprise;
5) copper selection: carrying out size mixing on the lead-zinc tailings obtained in the step 4) by adopting a copper selecting system liquid, controlling the concentration of the ore pulp to be 10 wt%, then adjusting the pH value of the ore pulp to be 5 by adopting sulfuric acid, adding 80g/t of ammonium nitrate black powder serving as a copper selecting agent and 60-80g/t of No. 2 oil serving as a foaming agent into the ore pulp, and carrying out primary roughing, secondary scavenging and secondary concentration operation to obtain copper concentrate containing 10 wt% of copper and second low-gold high-sulfur concentrate;
6) roasting to prepare acid and enrich gold, silver and iron to produce iron ore concentrate: feeding the second low-gold high-sulfur concentrate obtained in the step 5) into a fluidized bed roaster to roast in a weak oxygen reduction atmosphere, wherein the oxygen supply amount is 85-90% of the oxygen demand of oxidizing roasting, the roasting temperature is controlled at 700 ℃, roasting flue gas enters an acid making process to produce 98% or 105% sulfuric acid products, and iron concentrate with the sulfur content of 0.5%, the gold content of 2.5g/t, the silver content of 15g/t and the iron content of 62% is produced at the same time;
7) leaching gold and silver from iron ore concentrate by a wet method: mixing the iron ore concentrate obtained in the step 6) with filtrate obtained after filter pressing by a filter press, controlling the concentration of the ore pulp to be 30 wt%, then adjusting the pH value of the ore pulp to be 8 by adopting ammonia water, adding 2kg/t of environment-friendly medicament HBJ as a leaching agent, and after leaching for 6 hours, adsorbing the ore pulp by a carbon adsorption stirring tank, wherein the carbon density is 50kg/m3And adsorbing the ore pulp for 8 hours, then obtaining gold-loaded carbon and decarbonized ore pulp through a carbon separation sieve, obtaining finished gold from the gold-loaded carbon through a desorption electrolysis process, carrying out filter pressing on the decarbonized ore pulp through a filter press to obtain a filter cake and filtrate, returning the filtrate to a system for recycling, wherein the filter cake is iron concentrate with the sulfur content of 0.5 percent, the gold content of 0.5g/t, the silver content of 1g/t and the iron content of 62 percent, and selling the iron concentrate to a steel plant.
Example 2:
a method for comprehensively recovering multiple metal elements in gold concentrate through step separation comprises the following steps:
1) pre-flotation: the method comprises the steps of taking 30g/t gold-containing, 45g/t silver-containing, 0.3 wt% copper-containing, 0.3 wt% lead-containing, 0.3 wt% zinc-containing, 30 wt% sulfur-containing and 80% fineness-400 meshes as raw materials, carrying out liquid mixing with a flotation system, controlling the concentration of ore pulp to be 60 wt%, grinding the ore pulp until the fineness-400 meshes reach more than 92%, adding 1000g/t sodium metabisulfite and 1000g/t sodium hexametaphosphate as activating agents, 500g/t sodium xanthates and 500g/t 25# black chemicals as flotation collectors, 80g/t terpineol and 80g/t 2# oil as foaming agents, and carrying out primary roughing, secondary scavenging and secondary concentration processes to produce 50g/t gold-containing, 60g/t silver-containing, 0.5 wt% copper-containing, 0.5 wt% lead-containing, 0.5 wt% zinc-grade and 0.5 wt% zinc-grade position, High gold-sulfur concentrate with the sulfur grade of 48.3 wt% and low-sulfur tailings with the gold grade of 3.5g/t, the silver grade of 9g/t and the sulfur grade of 3.5 wt%;
2) alkaline leaching, cyaniding and gold extraction: the high gold and sulfur concentrate obtained in the step 1) is subjected to size mixing by adopting alkali leaching solution until the concentration of ore pulp is 35 wt%, lime is added to adjust the pH value of the ore pulp to be 10.5, and 40m is filled into the ore pulp3Air/hour, carrying out alkaline leaching reaction for 5 hours under the stirring condition of 40r/min, then carrying out filter pressing to obtain alkaline leaching liquid and an alkaline leaching filter cake, recycling the obtained alkaline leaching liquid, using first zinc powder to replace barren liquor for size mixing of the obtained alkaline leaching filter cake, controlling the concentration of ore pulp to be 30 wt%, carrying out cyaniding leaching, using sodium cyanide as a leaching agent, controlling the concentration of sodium cyanide to be 0.35%, leaching for 54 hours, fully washing and carrying out liquid-solid separation on the leached ore pulp to obtain gold-containing pregnant solution and first low-gold high-sulfur concentrate, and carrying out purification, deoxidation, zinc powder replacement and plate frame filter pressing on the gold-containing pregnant solution to obtain gold mud and first zinc powder replacement barren liquor;
3) leaching gold and silver from low-sulfur tailings by a wet method: mixing the low-sulfur tailings obtained in the step 1) by using second zinc powder to replace barren liquor, controlling the concentration of the ore pulp to be 30 wt%, adjusting the pH value of the ore pulp to be 9, adding 4kg/t of environment-friendly medicament HBJ serving as a leaching agent into the ore pulp, fully washing the ore pulp after leaching for 16 hours, performing liquid-solid separation to obtain gold-containing pregnant solution and low-sulfur tailings, and purifying, deoxidizing, replacing zinc powder and performing plate-and-frame filter pressing on the gold-containing pregnant solution to obtain gold mud and second zinc powder replacement barren liquor;
4) lead and zinc ore selection: carrying out liquid size mixing on the first low-gold high-sulfur concentrate generated in the step 2) by using a lead-selecting system, controlling the concentration of the ore pulp to be 20 wt%, then adjusting the pH value of the ore pulp to be 9 by using soda ash, adding 450g/t of copper sulfate serving as a zinc activating agent into the ore pulp, simultaneously adding 100g/t of ethidene nitrogen serving as a lead-zinc ore-selecting agent and 70g/t of 2# oil serving as a foaming agent into the ore pulp, carrying out once roughing, twice scavenging and twice selecting operations to obtain lead-zinc mixed concentrate containing 25% of lead, 25% of zinc and lead-zinc tailings, and selling the lead-zinc mixed concentrate to a lead-zinc smelting enterprise;
5) copper selection: carrying out size mixing on the lead-zinc tailings obtained in the step 4) by adopting a copper separation system liquid, controlling the concentration of the ore pulp to be 20 wt%, then adjusting the pH value of the ore pulp to be 6 by adopting acid waste liquid, adding 100g/t of butyl xanthate serving as a copper separation agent and 70g/t of 2# oil serving as a foaming agent into the ore pulp, and carrying out primary roughing, secondary scavenging and secondary concentration operation to obtain copper concentrate containing 15 wt% of copper and second low-gold high-sulfur concentrate;
6) roasting to prepare acid and enrich gold, silver and iron to produce iron ore concentrate: feeding the second low-gold high-sulfur concentrate obtained in the step 5) into a fluidized bed roaster to roast in a weak oxygen reduction atmosphere, wherein the oxygen supply amount is 85-90% of the oxygen demand of oxidizing roasting, the roasting temperature is controlled at 750 ℃, roasting flue gas enters an acid making process to produce 98% or 105% sulfuric acid products, and iron concentrate with the sulfur content of 0.4%, the gold content of 3.8g/t, the silver content of 35g/t and the iron content of 63.5% is produced;
7) leaching gold and silver from iron ore concentrate by a wet method: mixing the iron ore concentrate obtained in the step 6) with filtrate obtained after filter pressing by a filter press, controlling the concentration of the ore pulp to be 40 wt%, then adjusting the pH value of the ore pulp to be 9 by adopting ammonia water, adding 4kg/t of environment-friendly medicament HBJ as a leaching agent, and after leaching for 16 hours, adsorbing the ore pulp by a carbon adsorption stirring tank, wherein the carbon density is 75kg/m3And adsorbing the ore pulp for 12 hours, then obtaining gold-loaded carbon and decarbonized ore pulp through a carbon separation sieve, obtaining finished gold from the gold-loaded carbon through a desorption electrolysis process, carrying out filter pressing on the decarbonized ore pulp through a filter press to obtain a filter cake and filtrate, returning the filtrate to a system for recycling, wherein the filter cake is iron concentrate with the sulfur content of 0.4 percent, the gold content of 0.65g/t, the silver content of 3g/t and the iron content of 63.5 percent, and selling the iron concentrate to a steel mill.
Example 3:
a method for comprehensively recovering multiple metal elements in gold concentrate through step separation comprises the following steps:
1) pre-flotation: taking gold-containing gold concentrate 50g/t, silver-containing gold concentrate 80g/t, copper-containing gold concentrate 0.5 wt%, lead-containing gold concentrate 0.5 wt%, zinc-containing gold concentrate 0.5 wt%, sulfur-containing gold concentrate 40 wt% and low-sulfur gold concentrate 70% in fineness-400 meshes as raw materials, carrying out liquid mixing with a flotation system, controlling the concentration of the ore slurry to 70 wt%, grinding the ore slurry until the fineness-400 meshes reach 95%, adding sodium pyrosulfite, sodium hexametaphosphate and copper sulfate 1000g/t respectively as activating agents, butyl sodium xanthate, 25# black powder and butyl ammonium black powder 400g/t respectively as flotation collecting agents, terpenic oil and 2# oil 100g/t respectively as foaming agents, carrying out primary roughing and secondary scavenging and secondary concentration flotation processes to produce gold-containing gold-grade gold-containing gold-80 g/t, silver-grade-100 g/t, copper-0.8 wt%, lead-grade-0.8 wt%, zinc-grade-0.8 wt% and sulfur-containing gold-6.0 g/t, Low-sulfur tailings with silver grade of 15g/t and sulfur grade of 6 wt%;
2) alkaline leaching, cyaniding and gold extraction: the high gold and sulfur concentrate obtained in the step 1) is treatedThe pulp is adjusted by alkali immersion liquid until the pulp concentration is 50 wt%, lime is added to adjust the pH value of the pulp to be 12, and 50m is charged into the pulp3Air per hour is subjected to alkaline leaching reaction for 6 hours under the stirring condition of 50r/min, then, alkaline leaching liquid and alkaline leaching filter cake are obtained through pressure filtration, the obtained alkaline leaching liquid is recycled, the obtained alkaline leaching filter cake is subjected to size mixing by adopting first zinc powder replacement barren solution, the concentration of ore pulp is controlled to be 50 wt%, then, cyaniding leaching is carried out, sodium cyanide is used as a leaching agent, the concentration of sodium cyanide is controlled to be 0.60%, leaching is carried out for 72 hours, the leached ore pulp is subjected to full washing and liquid-solid separation to obtain gold-containing pregnant solution and first low-gold high-sulfur concentrate, and the gold-containing pregnant solution is subjected to purification, deoxidation, zinc powder replacement and plate frame pressure filtration to obtain gold mud and first zinc powder replacement barren solution;
3) leaching gold and silver from low-sulfur tailings by a wet method: mixing the low-sulfur tailings obtained in the step 1) with a second zinc powder displacement barren solution, controlling the concentration of the ore pulp to be 50 wt%, adjusting the pH value of the ore pulp to be 10 by using ammonia water, adding 6kg/t of environment-friendly medicament HBJ serving as a leaching agent into the ore pulp, fully washing the ore pulp after leaching for 24 hours, performing liquid-solid separation to obtain gold-containing pregnant solution and low-sulfur tailings, and purifying, deoxidizing, replacing zinc powder, and performing plate-and-frame filter pressing on the gold-containing pregnant solution to obtain gold mud and a second zinc powder displacement barren solution;
4) lead and zinc ore selection: carrying out liquid size mixing on the first low-gold high-sulfur concentrate generated in the step 2) by using a lead-selecting system, controlling the concentration of the ore pulp to be 30 wt%, then adjusting the pH value of the ore pulp to be 10 by using soda ash, adding 600g/t of copper sulfate as a zinc activator, simultaneously adding 60g/t of butyl xanthate and ethidium chloride as lead-zinc ore-dressing agents, adding 80g/t of No. 2 oil as a foaming agent, carrying out once rough concentration and twice scavenging and twice concentration operations to obtain lead-zinc mixed concentrate containing 30% of lead and 30% of zinc and lead-zinc tailings, and selling the lead-zinc mixed concentrate to a lead-zinc smelting enterprise;
5) copper selection: carrying out size mixing on the lead-zinc tailings obtained in the step 4) by adopting a copper separation system liquid, controlling the concentration of the ore pulp to be 30 wt%, then adjusting the pH value of the ore pulp to be 7 by adopting sulfuric acid, adding 60g/t of each of butyl xanthate and butyl ammonium nigride as a copper separation agent and 80g/t of No. 2 oil as a foaming agent, and carrying out primary roughing, secondary scavenging and secondary concentration operation to obtain copper concentrate containing 20 wt% of copper and second low-gold high-sulfur concentrate;
6) roasting to prepare acid and enrich gold, silver and iron to produce iron ore concentrate: feeding the second low-gold high-sulfur concentrate obtained in the step 5) into a fluidized bed roaster to roast in a weak oxygen reduction atmosphere, wherein the oxygen supply amount is 85-90% of the oxygen demand of oxidizing roasting, the roasting temperature is controlled at 800 ℃, roasting flue gas enters an acid making process to produce 98% or 105% sulfuric acid products, and iron concentrate with 0.3% of sulfur content, 5.0g/t of gold, 55g/t of silver and 65% of iron is produced;
7) leaching gold and silver from iron ore concentrate by a wet method: mixing the iron ore concentrate obtained in the step 6) with filtrate obtained after filter pressing by a filter press, controlling the concentration of the ore pulp to be 50 wt%, then adjusting the pH value of the ore pulp to be 10 by adopting ammonia water, adding 6kg/t of environment-friendly medicament HBJ as a leaching agent, and after leaching for 24 hours, adsorbing the ore pulp by a carbon adsorption stirring tank, wherein the carbon density is 100kg/m3And (2) obtaining gold-loaded carbon and decarburized ore pulp by passing the ore pulp through a carbon separation sieve, obtaining finished gold from the gold-loaded carbon through a desorption electrolysis process, carrying out filter pressing on the decarburized ore pulp through a filter press to obtain a filter cake and filtrate, returning the filtrate to a system for recycling, wherein the filter cake is iron ore concentrate with the sulfur content of 0.3%, the gold content of 0.8g/t, the silver content of 5g/t and the iron content of 65%, and selling the iron ore concentrate to a steel plant.
Example 4:
a method for comprehensively recovering multiple metal elements in gold concentrate through step separation comprises the following steps:
1) pre-flotation: taking 30g/t gold-containing gold, 45g/t silver-containing, 0.3 wt% copper-containing, 0.3 wt% lead-containing, 0.3 wt% zinc-containing, 30 wt% sulfur-containing and 80% fineness-400 meshes as raw materials, carrying out liquid mixing with a flotation system, controlling the concentration of the ore pulp to be 60 wt%, grinding the ore pulp until the fineness-400 meshes reach more than 92%, adding 2000g/t sodium hexametaphosphate as an activating agent, 100g/t 25# black drug as a flotation collecting agent, 150g/t terpineol oil as a foaming agent, carrying out primary roughing, secondary scavenging and secondary concentration flotation processes to produce 50g/t gold-containing, 60g/t silver-containing, 0.5 wt% copper-containing, 0.5 wt% lead-containing, 0.5 wt% zinc-containing and 48.3 wt% sulfur-containing high gold-sulfur concentrate and 3.5g/t gold-containing, 9g/t silver-containing, Low-sulfur tailings with 3.5 wt% sulfur grade;
2) alkaline leaching of cyanogenGold extraction: the high gold and sulfur concentrate obtained in the step 1) is subjected to size mixing by adopting alkali leaching solution until the concentration of ore pulp is 35 wt%, lime is added to adjust the pH value of the ore pulp to be 10.5, and 40m is filled into the ore pulp3Air/hour, carrying out alkaline leaching reaction for 5 hours under the stirring condition of 40r/min, then carrying out filter pressing to obtain alkaline leaching liquid and an alkaline leaching filter cake, recycling the obtained alkaline leaching liquid, using first zinc powder to replace barren liquor for size mixing of the obtained alkaline leaching filter cake, controlling the concentration of ore pulp to be 30 wt%, carrying out cyaniding leaching, using sodium cyanide as a leaching agent, controlling the concentration of sodium cyanide to be 0.35%, leaching for 54 hours, fully washing and carrying out liquid-solid separation on the leached ore pulp to obtain gold-containing pregnant solution and first low-gold high-sulfur concentrate, and carrying out purification, deoxidation, zinc powder replacement and plate frame filter pressing on the gold-containing pregnant solution to obtain gold mud and first zinc powder replacement barren liquor;
3) leaching gold and silver from low-sulfur tailings by a wet method: mixing the low-sulfur tailings obtained in the step 1) by using second zinc powder to replace barren liquor, controlling the concentration of the ore pulp to be 30 wt%, adjusting the pH value of the ore pulp to be 9, adding 4kg/t of environment-friendly medicament HBJ serving as a leaching agent into the ore pulp, fully washing the ore pulp after leaching for 16 hours, performing liquid-solid separation to obtain gold-containing pregnant solution and low-sulfur tailings, and purifying, deoxidizing, replacing zinc powder and performing plate-and-frame filter pressing on the gold-containing pregnant solution to obtain gold mud and second zinc powder replacement barren liquor;
4) lead and zinc ore selection: carrying out liquid size mixing on the first low-gold high-sulfur concentrate generated in the step 2) by using a lead-selecting system, controlling the concentration of the ore pulp to be 20 wt%, then adjusting the pH value of the ore pulp to be 9 by using soda ash, adding 450g/t of copper sulfate serving as a zinc activating agent into the ore pulp, simultaneously adding 100g/t of ethidene nitrogen serving as a lead-zinc ore-selecting agent and 70g/t of 2# oil serving as a foaming agent into the ore pulp, carrying out once roughing, twice scavenging and twice selecting operations to obtain lead-zinc mixed concentrate containing 25% of lead, 25% of zinc and lead-zinc tailings, and selling the lead-zinc mixed concentrate to a lead-zinc smelting enterprise;
5) copper selection: carrying out size mixing on the lead-zinc tailings obtained in the step 4) by adopting a copper separation system liquid, controlling the concentration of the ore pulp to be 20 wt%, then adjusting the pH value of the ore pulp to be 6 by adopting acid waste liquid, adding 100g/t of butyl xanthate serving as a copper separation agent and 70g/t of 2# oil serving as a foaming agent into the ore pulp, and carrying out primary roughing, secondary scavenging and secondary concentration operation to obtain copper concentrate containing 15 wt% of copper and second low-gold high-sulfur concentrate;
6) roasting to prepare acid and enrich gold, silver and iron to produce iron ore concentrate: feeding the second low-gold high-sulfur concentrate obtained in the step 5) into a fluidized bed roaster to roast in a weak oxygen reduction atmosphere, wherein the oxygen supply amount is 85-90% of the oxygen demand of oxidizing roasting, the roasting temperature is controlled at 750 ℃, roasting flue gas enters an acid making process to produce 98% or 105% sulfuric acid products, and iron concentrate with the sulfur content of 0.4%, the gold content of 3.8g/t, the silver content of 35g/t and the iron content of 63.5% is produced;
7) leaching gold and silver from iron ore concentrate by a wet method: mixing the iron ore concentrate obtained in the step 6) with filtrate obtained after filter pressing by a filter press, controlling the concentration of the ore pulp to be 40 wt%, then adjusting the pH value of the ore pulp to be 9 by adopting ammonia water, adding 4kg/t of environment-friendly medicament HBJ as a leaching agent, and after leaching for 16 hours, adsorbing the ore pulp by a carbon adsorption stirring tank, wherein the carbon density is 75kg/m3And adsorbing the ore pulp for 12 hours, then obtaining gold-loaded carbon and decarbonized ore pulp through a carbon separation sieve, obtaining finished gold from the gold-loaded carbon through a desorption electrolysis process, carrying out filter pressing on the decarbonized ore pulp through a filter press to obtain a filter cake and filtrate, returning the filtrate to a system for recycling, wherein the filter cake is iron concentrate with the sulfur content of 0.4 percent, the gold content of 0.65g/t, the silver content of 3g/t and the iron content of 63.5 percent, and selling the iron concentrate to a steel mill.
Example 5:
a method for comprehensively recovering multiple metal elements in gold concentrate through step separation comprises the following steps:
1) pre-flotation: using 30g/t gold-containing, 45g/t silver-containing, 0.3 wt% copper-containing, 0.3 wt% lead-containing, 0.3 wt% zinc-containing, 30 wt% sulfur-containing and 80% fineness-400 mesh as raw materials, mixing the raw materials with flotation system liquid, controlling the concentration of the ore pulp to be 60 wt%, grinding the ore pulp until the fineness-400 mesh reaches more than 92%, adding 2500g/t copper sulfate as an activating agent, 1000g/t butylamine black as a flotation collecting agent, 80g/t pine oil and 80g/t 2# oil as foaming agents, performing primary roughing and secondary scavenging and secondary concentration flotation processes to produce high gold sulfur concentrate with a gold-containing grade of 50g/t, 60g/t silver, 0.5 wt% copper, 0.5 wt% lead, 0.5 wt% zinc grade and 48.3 wt% sulfur grade and gold-containing grade of 3.5g/t, Low-sulfur tailings with 9g/t silver grade and 3.5 wt% sulfur grade;
2) alkaline leaching, cyaniding and gold extraction: the high gold and sulfur concentrate obtained in the step 1) is subjected to size mixing by adopting alkali leaching solution until the concentration of ore pulp is 35 wt%, lime is added to adjust the pH value of the ore pulp to be 10.5, and 40m is filled into the ore pulp3Air/hour, carrying out alkaline leaching reaction for 5 hours under the stirring condition of 40r/min, then carrying out filter pressing to obtain alkaline leaching liquid and an alkaline leaching filter cake, recycling the obtained alkaline leaching liquid, using first zinc powder to replace barren liquor for size mixing of the obtained alkaline leaching filter cake, controlling the concentration of ore pulp to be 30 wt%, carrying out cyaniding leaching, using sodium cyanide as a leaching agent, controlling the concentration of sodium cyanide to be 0.35%, leaching for 54 hours, fully washing and carrying out liquid-solid separation on the leached ore pulp to obtain gold-containing pregnant solution and first low-gold high-sulfur concentrate, and carrying out purification, deoxidation, zinc powder replacement and plate frame filter pressing on the gold-containing pregnant solution to obtain gold mud and first zinc powder replacement barren liquor;
3) leaching gold and silver from low-sulfur tailings by a wet method: mixing the low-sulfur tailings obtained in the step 1) by using second zinc powder to replace barren liquor, controlling the concentration of the ore pulp to be 30 wt%, adjusting the pH value of the ore pulp to be 9, adding 4kg/t of environment-friendly medicament HBJ serving as a leaching agent into the ore pulp, fully washing the ore pulp after leaching for 16 hours, performing liquid-solid separation to obtain gold-containing pregnant solution and low-sulfur tailings, and purifying, deoxidizing, replacing zinc powder and performing plate-and-frame filter pressing on the gold-containing pregnant solution to obtain gold mud and second zinc powder replacement barren liquor;
4) lead and zinc ore selection: carrying out liquid size mixing on the first low-gold high-sulfur concentrate generated in the step 2) by using a lead-selecting system, controlling the concentration of the ore pulp to be 20 wt%, then adjusting the pH value of the ore pulp to be 9 by using soda ash, adding 450g/t of copper sulfate serving as a zinc activating agent into the ore pulp, simultaneously adding 100g/t of ethidene nitrogen serving as a lead-zinc ore-selecting agent and 70g/t of 2# oil serving as a foaming agent into the ore pulp, carrying out once roughing, twice scavenging and twice selecting operations to obtain lead-zinc mixed concentrate containing 25% of lead, 25% of zinc and lead-zinc tailings, and selling the lead-zinc mixed concentrate to a lead-zinc smelting enterprise;
5) copper selection: carrying out size mixing on the lead-zinc tailings obtained in the step 4) by adopting a copper separation system liquid, controlling the concentration of the ore pulp to be 20 wt%, then adjusting the pH value of the ore pulp to be 6 by adopting acid waste liquid, adding 100g/t of butyl xanthate serving as a copper separation agent and 70g/t of 2# oil serving as a foaming agent into the ore pulp, and carrying out primary roughing, secondary scavenging and secondary concentration operation to obtain copper concentrate containing 15 wt% of copper and second low-gold high-sulfur concentrate;
6) roasting to prepare acid and enrich gold, silver and iron to produce iron ore concentrate: feeding the second low-gold high-sulfur concentrate obtained in the step 5) into a fluidized bed roaster to roast in a weak oxygen reduction atmosphere, wherein the oxygen supply amount is 85-90% of the oxygen demand of oxidizing roasting, the roasting temperature is controlled at 750 ℃, roasting flue gas enters an acid making process to produce 98% or 105% sulfuric acid products, and iron concentrate with the sulfur content of 0.4%, the gold content of 3.8g/t, the silver content of 35g/t and the iron content of 63.5% is produced;
7) leaching gold and silver from iron ore concentrate by a wet method: mixing the iron ore concentrate obtained in the step 6) with filtrate obtained after filter pressing by a filter press, controlling the concentration of the ore pulp to be 40 wt%, then adjusting the pH value of the ore pulp to be 9 by adopting ammonia water, adding 4kg/t of environment-friendly medicament HBJ as a leaching agent, and after leaching for 16 hours, adsorbing the ore pulp by a carbon adsorption stirring tank, wherein the carbon density is 75kg/m3And adsorbing the ore pulp for 12 hours, then obtaining gold-loaded carbon and decarbonized ore pulp through a carbon separation sieve, obtaining finished gold from the gold-loaded carbon through a desorption electrolysis process, carrying out filter pressing on the decarbonized ore pulp through a filter press to obtain a filter cake and filtrate, returning the filtrate to a system for recycling, wherein the filter cake is iron concentrate with the sulfur content of 0.4 percent, the gold content of 0.65g/t, the silver content of 3g/t and the iron content of 63.5 percent, and selling the iron concentrate to a steel mill.
The above description is only for the purpose of illustrating the preferred embodiments of the present invention and is not to be construed as limiting the invention, and any modifications, equivalents, improvements and the like that fall within the spirit and principle of the present invention are intended to be included therein.
Claims (9)
1. A method for comprehensively recovering multiple metal elements in gold concentrate through step separation is characterized by comprising the following steps:
1) pre-flotation: taking gold-containing 10-50g/t, silver-containing 10-80g/t, copper-containing 0.1-0.5 wt%, lead-containing 0.1-0.5 wt%, zinc-containing 0.1-0.5 wt%, sulfur-containing 20-40 wt% and low-sulfur gold concentrate with fineness of-400 meshes lower than 90% as raw materials, carrying out liquid mixing with a flotation system, grinding ore pulp until the fineness of-400 meshes reaches more than 90%, adding 1000-3000g/t of activating agent, 800-1200g/t of flotation collecting agent and 100-200g/t of foaming agent into the ore pulp, wherein the activating agent is one or a mixture of sodium pyrosulfite, sodium hexametaphosphate and copper sulfate, the flotation collecting agent is one or a mixture of sodium xanthates, 25# melanoides and butylammonium melanoides, and the foaming agent is one or a mixture of pine alcohol oil and 2# oil, through the processes of primary roughing, secondary scavenging and secondary concentration flotation, high gold-sulfur concentrate with gold grade of 20-80g/t, silver grade of 20-100g/t, copper grade of 0.2-0.8 wt%, lead grade of 0.2-0.8 wt%, zinc grade of 0.2-0.8 wt% and sulfur grade of not less than 48 wt% and low-sulfur tailings with gold grade of 1.0-6.0g/t, silver grade of 3-15g/t and sulfur grade of 1-6 wt% are produced;
2) alkaline leaching, cyaniding and gold extraction: the high gold and sulfur concentrate obtained in the step 1) is subjected to size mixing by adopting alkali immersion liquid, the pH value of ore pulp is adjusted to 9-12, and 30-50m is filled into the ore pulp3Air per hour is subjected to alkaline leaching reaction under the stirring condition, then alkaline leaching liquid and alkaline leaching filter cake are obtained through pressure filtration, the obtained alkaline leaching liquid is recycled, the obtained alkaline leaching filter cake is subjected to size mixing by adopting a first zinc powder displacement barren solution, then cyaniding leaching is carried out, the concentration of cyanide is controlled to be 0.10-0.60%, leached ore pulp is fully washed and subjected to liquid-solid separation to obtain gold-containing pregnant solution and a first low-gold high-sulfur concentrate, and the gold-containing pregnant solution is subjected to purification, deoxidation, zinc powder displacement and plate-and-frame pressure filtration to obtain gold mud and a first zinc powder displacement barren solution;
3) leaching gold and silver from low-sulfur tailings by a wet method: mixing the low-sulfur tailings obtained in the step 1) with a second zinc powder displacement barren solution, adjusting the pH value of ore pulp to 8-10, adding 2-6kg/t of environment-friendly medicament HBJ serving as a leaching agent into the ore pulp, fully washing the leached ore pulp and performing liquid-solid separation to obtain gold-containing pregnant solution and low-sulfur tailings, and purifying, deoxidizing, zinc powder displacement and plate-and-frame filter pressing the gold-containing pregnant solution to obtain gold mud and a second zinc powder displacement barren solution;
4) lead and zinc ore selection: mixing the first low-gold high-sulfur concentrate generated in the step 2) by adopting a lead concentration system, then adjusting the pH value of the ore pulp to be 8-10, adding 300-600g/t of copper sulfate as a zinc activator, simultaneously adding 80-120g/t of lead-zinc ore dressing agent and 60-80g/t of foaming agent, wherein the lead-zinc ore dressing agent is one or a mixture of butyl xanthate and ethidium nitrate, the foaming agent is No. 2 oil, obtaining 20-30 wt% of lead-zinc mixed concentrate and 20-30 wt% of zinc by performing primary rough concentration and secondary scavenging and secondary concentration, and selling the lead-zinc mixed concentrate to a lead-zinc smelting enterprise;
5) copper selection: carrying out size mixing on the lead-zinc tailings obtained in the step 4) by adopting a copper selecting system liquid, then adjusting the pH value of ore pulp to 5-7, adding 80-120g/t of copper selecting agent and 60-80g/t of foaming agent into the ore pulp, wherein the copper selecting agent is one or a mixture of butyl xanthate and butyl ammonium nigride, the foaming agent is 2# oil, and carrying out once roughing, twice scavenging and twice fine concentration operation to obtain copper concentrate containing 10-20 wt% of copper and second low-gold high-sulfur concentrate;
6) roasting to prepare acid and enrich gold, silver and iron to produce iron ore concentrate: feeding the second low-gold high-sulfur concentrate obtained in the step 5) into a fluidized bed roaster for roasting in a weak oxygen reduction atmosphere, wherein the oxygen supply amount is 85-90% of the oxygen demand of oxidizing roasting, the roasting temperature is controlled at 800 ℃ with 700-;
7) leaching gold and silver from iron ore concentrate by a wet method: mixing the iron ore concentrate obtained in the step 6) with the filtrate obtained after filter pressing by a filter press, then adjusting the pH value of the ore slurry to 8-10, adding 2-6kg/t of environment-friendly medicament HBJ as a leaching agent, adsorbing the leached ore slurry by a carbon adsorption stirring tank, wherein the carbon density is 50-100kg/m3Ore pulp, obtaining gold-loaded carbon and decarbonized ore pulp through a carbon separation sieve, obtaining finished gold from the gold-loaded carbon through a desorption electrolysis process, obtaining filter cakes and filtrate from the decarbonized ore pulp through a filter press, returning the filtrate to a system for recycling, wherein the filter cakes are iron ore concentrates with the sulfur content of less than or equal to 0.5 wt%, the gold content of 0.5-0.8g/t, the silver content of 1-5g/t and the iron content of 62-65 wt%, and selling the iron ore concentrates to a steel mill;
wherein, the step 2) and the step 3) are not in sequence.
2. The method according to claim 1, characterized in that the pulp concentration in step 1) is controlled to 50-70 wt%.
3. The method as claimed in claim 1, wherein the concentration of the pulp is controlled by adjusting the alkaline leaching solution to 20-50 wt% in step 2), and the pulp is adjusted to 10-50 wt% by using the first zinc powder to replace the barren solution.
4. The method according to claim 1 or 3, wherein the stirring speed in the step 2) is 30-50r/min, and the alkaline leaching reaction time is 4-6 hours.
5. The method as claimed in claim 1, wherein the pulp concentration is controlled to 10-50 wt% in step 3).
6. The method as claimed in claim 1, wherein the leaching time in step 3) is 8-24 hours.
7. The method according to claim 1, characterized in that the pulp concentration is controlled to 10-30 wt% in step 4) and step 5).
8. The method as claimed in claim 1, wherein the pulp concentration is controlled to 30-50 wt% in step 7).
9. The method as claimed in claim 1, wherein the leaching time in step 7) is 8-24 hours and the adsorption process in step 7) is 8-16 hours.
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| CN110306066B (en) * | 2019-07-22 | 2021-07-20 | 山东国大黄金股份有限公司 | Multistage gold extraction method for refractory arsenic-containing gold tailings |
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Denomination of invention: A method for step separation and comprehensive recovery of multiple metal elements in gold concentrate Effective date of registration: 20211202 Granted publication date: 20200512 Pledgee: Zhaoyuan Huatang Trading Co.,Ltd. Pledgor: SHANDONG GUODA GOLD Co.,Ltd. Registration number: Y2021980013873 |
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| PC01 | Cancellation of the registration of the contract for pledge of patent right |
Date of cancellation: 20220622 Granted publication date: 20200512 Pledgee: Zhaoyuan Huatang Trading Co.,Ltd. Pledgor: SHANDONG GUODA GOLD Co.,Ltd. Registration number: Y2021980013873 |