CN111100988A - Method for reducing and utilizing multi-metal tailings - Google Patents
Method for reducing and utilizing multi-metal tailings Download PDFInfo
- Publication number
- CN111100988A CN111100988A CN201911285182.0A CN201911285182A CN111100988A CN 111100988 A CN111100988 A CN 111100988A CN 201911285182 A CN201911285182 A CN 201911285182A CN 111100988 A CN111100988 A CN 111100988A
- Authority
- CN
- China
- Prior art keywords
- solution
- filtrate
- tailings
- leaching
- sodium
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Granted
Links
- 238000000034 method Methods 0.000 title claims abstract description 48
- 229910052751 metal Inorganic materials 0.000 title claims abstract description 41
- 239000002184 metal Substances 0.000 title claims abstract description 34
- CDBYLPFSWZWCQE-UHFFFAOYSA-L Sodium Carbonate Chemical compound [Na+].[Na+].[O-]C([O-])=O CDBYLPFSWZWCQE-UHFFFAOYSA-L 0.000 claims abstract description 40
- XLOMVQKBTHCTTD-UHFFFAOYSA-N Zinc monoxide Chemical compound [Zn]=O XLOMVQKBTHCTTD-UHFFFAOYSA-N 0.000 claims abstract description 39
- 238000002386 leaching Methods 0.000 claims abstract description 36
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims abstract description 35
- 239000010949 copper Substances 0.000 claims abstract description 34
- BIGPRXCJEDHCLP-UHFFFAOYSA-N ammonium bisulfate Chemical compound [NH4+].OS([O-])(=O)=O BIGPRXCJEDHCLP-UHFFFAOYSA-N 0.000 claims abstract description 32
- 239000011701 zinc Substances 0.000 claims abstract description 26
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 claims abstract description 25
- 229910052802 copper Inorganic materials 0.000 claims abstract description 25
- 238000001027 hydrothermal synthesis Methods 0.000 claims abstract description 25
- 229910052710 silicon Inorganic materials 0.000 claims abstract description 25
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims abstract description 24
- XUIMIQQOPSSXEZ-UHFFFAOYSA-N Silicon Chemical compound [Si] XUIMIQQOPSSXEZ-UHFFFAOYSA-N 0.000 claims abstract description 24
- 239000010703 silicon Substances 0.000 claims abstract description 24
- 150000002500 ions Chemical class 0.000 claims abstract description 20
- 229910000029 sodium carbonate Inorganic materials 0.000 claims abstract description 20
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 claims abstract description 19
- 229910052782 aluminium Inorganic materials 0.000 claims abstract description 19
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 claims abstract description 19
- 229910052725 zinc Inorganic materials 0.000 claims abstract description 17
- 239000011787 zinc oxide Substances 0.000 claims abstract description 17
- 238000001556 precipitation Methods 0.000 claims abstract description 16
- 238000002156 mixing Methods 0.000 claims abstract description 14
- 239000002253 acid Substances 0.000 claims abstract description 11
- 239000000741 silica gel Substances 0.000 claims abstract description 11
- 229910002027 silica gel Inorganic materials 0.000 claims abstract description 11
- 239000006229 carbon black Substances 0.000 claims abstract description 9
- 239000011941 photocatalyst Substances 0.000 claims abstract description 6
- 238000000605 extraction Methods 0.000 claims abstract description 5
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 108
- 239000000243 solution Substances 0.000 claims description 82
- 239000000706 filtrate Substances 0.000 claims description 49
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims description 47
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 41
- 239000002893 slag Substances 0.000 claims description 41
- 239000007788 liquid Substances 0.000 claims description 39
- 235000011121 sodium hydroxide Nutrition 0.000 claims description 36
- 239000000843 powder Substances 0.000 claims description 34
- NTHWMYGWWRZVTN-UHFFFAOYSA-N sodium silicate Chemical compound [Na+].[Na+].[O-][Si]([O-])=O NTHWMYGWWRZVTN-UHFFFAOYSA-N 0.000 claims description 33
- 238000000926 separation method Methods 0.000 claims description 27
- 229910052500 inorganic mineral Inorganic materials 0.000 claims description 26
- 235000010755 mineral Nutrition 0.000 claims description 26
- 239000011707 mineral Substances 0.000 claims description 26
- 239000002244 precipitate Substances 0.000 claims description 24
- 229910052742 iron Inorganic materials 0.000 claims description 22
- 239000007791 liquid phase Substances 0.000 claims description 21
- 238000001914 filtration Methods 0.000 claims description 19
- PIJPYDMVFNTHIP-UHFFFAOYSA-L lead sulfate Chemical compound [PbH4+2].[O-]S([O-])(=O)=O PIJPYDMVFNTHIP-UHFFFAOYSA-L 0.000 claims description 19
- 238000006243 chemical reaction Methods 0.000 claims description 18
- 239000004115 Sodium Silicate Substances 0.000 claims description 17
- 229910052911 sodium silicate Inorganic materials 0.000 claims description 17
- 235000019353 potassium silicate Nutrition 0.000 claims description 16
- 239000007787 solid Substances 0.000 claims description 16
- BFNBIHQBYMNNAN-UHFFFAOYSA-N ammonium sulfate Chemical compound N.N.OS(O)(=O)=O BFNBIHQBYMNNAN-UHFFFAOYSA-N 0.000 claims description 14
- 229910052921 ammonium sulfate Inorganic materials 0.000 claims description 14
- 235000011130 ammonium sulphate Nutrition 0.000 claims description 14
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 claims description 12
- FAPWRFPIFSIZLT-UHFFFAOYSA-M Sodium chloride Chemical compound [Na+].[Cl-] FAPWRFPIFSIZLT-UHFFFAOYSA-M 0.000 claims description 12
- VTYYLEPIZMXCLO-UHFFFAOYSA-L Calcium carbonate Chemical compound [Ca+2].[O-]C([O-])=O VTYYLEPIZMXCLO-UHFFFAOYSA-L 0.000 claims description 10
- PMZURENOXWZQFD-UHFFFAOYSA-L Sodium Sulfate Chemical compound [Na+].[Na+].[O-]S([O-])(=O)=O PMZURENOXWZQFD-UHFFFAOYSA-L 0.000 claims description 10
- 238000000227 grinding Methods 0.000 claims description 10
- 239000000047 product Substances 0.000 claims description 10
- 235000012239 silicon dioxide Nutrition 0.000 claims description 10
- 229910052938 sodium sulfate Inorganic materials 0.000 claims description 10
- 235000011152 sodium sulphate Nutrition 0.000 claims description 10
- 238000005406 washing Methods 0.000 claims description 10
- 239000000499 gel Substances 0.000 claims description 9
- RMAQACBXLXPBSY-UHFFFAOYSA-N silicic acid Chemical compound O[Si](O)(O)O RMAQACBXLXPBSY-UHFFFAOYSA-N 0.000 claims description 9
- CWYNVVGOOAEACU-UHFFFAOYSA-N Fe2+ Chemical compound [Fe+2] CWYNVVGOOAEACU-UHFFFAOYSA-N 0.000 claims description 8
- 229920002538 Polyethylene Glycol 20000 Polymers 0.000 claims description 8
- 239000007789 gas Substances 0.000 claims description 8
- VTLYFUHAOXGGBS-UHFFFAOYSA-N Fe3+ Chemical compound [Fe+3] VTLYFUHAOXGGBS-UHFFFAOYSA-N 0.000 claims description 7
- 229910001447 ferric ion Inorganic materials 0.000 claims description 7
- MHAJPDPJQMAIIY-UHFFFAOYSA-N Hydrogen peroxide Chemical group OO MHAJPDPJQMAIIY-UHFFFAOYSA-N 0.000 claims description 6
- CBENFWSGALASAD-UHFFFAOYSA-N Ozone Chemical compound [O-][O+]=O CBENFWSGALASAD-UHFFFAOYSA-N 0.000 claims description 6
- 238000001035 drying Methods 0.000 claims description 6
- 239000003792 electrolyte Substances 0.000 claims description 6
- 229910052935 jarosite Inorganic materials 0.000 claims description 6
- 239000011780 sodium chloride Substances 0.000 claims description 6
- HFQQZARZPUDIFP-UHFFFAOYSA-M sodium;2-dodecylbenzenesulfonate Chemical compound [Na+].CCCCCCCCCCCCC1=CC=CC=C1S([O-])(=O)=O HFQQZARZPUDIFP-UHFFFAOYSA-M 0.000 claims description 6
- 238000003756 stirring Methods 0.000 claims description 6
- QGZKDVFQNNGYKY-UHFFFAOYSA-O Ammonium Chemical compound [NH4+] QGZKDVFQNNGYKY-UHFFFAOYSA-O 0.000 claims description 5
- ATRRKUHOCOJYRX-UHFFFAOYSA-N Ammonium bicarbonate Chemical compound [NH4+].OC([O-])=O ATRRKUHOCOJYRX-UHFFFAOYSA-N 0.000 claims description 5
- 229910000013 Ammonium bicarbonate Inorganic materials 0.000 claims description 5
- JPVYNHNXODAKFH-UHFFFAOYSA-N Cu2+ Chemical compound [Cu+2] JPVYNHNXODAKFH-UHFFFAOYSA-N 0.000 claims description 5
- 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 5
- WNROFYMDJYEPJX-UHFFFAOYSA-K aluminium hydroxide Chemical compound [OH-].[OH-].[OH-].[Al+3] WNROFYMDJYEPJX-UHFFFAOYSA-K 0.000 claims description 5
- 235000012538 ammonium bicarbonate Nutrition 0.000 claims description 5
- 239000001099 ammonium carbonate Substances 0.000 claims description 5
- 239000007864 aqueous solution Substances 0.000 claims description 5
- 229910000019 calcium carbonate Inorganic materials 0.000 claims description 5
- 229910001431 copper ion Inorganic materials 0.000 claims description 5
- 229910021645 metal ion Inorganic materials 0.000 claims description 5
- 150000003839 salts Chemical class 0.000 claims description 5
- 239000011734 sodium Substances 0.000 claims description 5
- 229910052708 sodium Inorganic materials 0.000 claims description 5
- -1 zincate ions Chemical class 0.000 claims description 5
- 235000008733 Citrus aurantifolia Nutrition 0.000 claims description 4
- 235000011941 Tilia x europaea Nutrition 0.000 claims description 4
- 238000010000 carbonizing Methods 0.000 claims description 4
- 239000000839 emulsion Substances 0.000 claims description 4
- 239000004571 lime Substances 0.000 claims description 4
- 239000000203 mixture Substances 0.000 claims description 3
- 238000012545 processing Methods 0.000 claims description 3
- 230000035484 reaction time Effects 0.000 claims description 3
- 238000010438 heat treatment Methods 0.000 claims description 2
- 230000001590 oxidative effect Effects 0.000 claims description 2
- 239000002245 particle Substances 0.000 claims description 2
- 238000010828 elution Methods 0.000 claims 4
- 238000000151 deposition Methods 0.000 claims 3
- 230000008021 deposition Effects 0.000 claims 2
- 229910001385 heavy metal Inorganic materials 0.000 abstract description 9
- 238000011084 recovery Methods 0.000 abstract description 7
- 239000003513 alkali Substances 0.000 abstract description 3
- 238000004064 recycling Methods 0.000 abstract description 3
- 239000002910 solid waste Substances 0.000 abstract description 3
- 239000000284 extract Substances 0.000 abstract description 2
- 231100000086 high toxicity Toxicity 0.000 abstract description 2
- 229910001804 ammoniojarosite Inorganic materials 0.000 abstract 1
- 150000003751 zinc Chemical class 0.000 abstract 1
- 239000011133 lead Substances 0.000 description 10
- QGZKDVFQNNGYKY-UHFFFAOYSA-N Ammonia Chemical compound N QGZKDVFQNNGYKY-UHFFFAOYSA-N 0.000 description 8
- 238000004090 dissolution Methods 0.000 description 8
- QAOWNCQODCNURD-UHFFFAOYSA-L Sulfate Chemical compound [O-]S([O-])(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-L 0.000 description 5
- 229910001570 bauxite Inorganic materials 0.000 description 5
- 238000005065 mining Methods 0.000 description 5
- 238000012360 testing method Methods 0.000 description 4
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 description 3
- 229910052793 cadmium Inorganic materials 0.000 description 3
- BDOSMKKIYDKNTQ-UHFFFAOYSA-N cadmium atom Chemical compound [Cd] BDOSMKKIYDKNTQ-UHFFFAOYSA-N 0.000 description 3
- 239000002994 raw material Substances 0.000 description 3
- 229910052709 silver Inorganic materials 0.000 description 3
- 239000004332 silver Substances 0.000 description 3
- ONIOAEVPMYCHKX-UHFFFAOYSA-N carbonic acid;zinc Chemical compound [Zn].OC(O)=O ONIOAEVPMYCHKX-UHFFFAOYSA-N 0.000 description 2
- 230000000694 effects Effects 0.000 description 2
- 238000005516 engineering process Methods 0.000 description 2
- 230000007613 environmental effect Effects 0.000 description 2
- 150000002739 metals Chemical class 0.000 description 2
- 239000002366 mineral element Substances 0.000 description 2
- 238000002360 preparation method Methods 0.000 description 2
- 239000000377 silicon dioxide Substances 0.000 description 2
- 239000002689 soil Substances 0.000 description 2
- 239000001117 sulphuric acid Substances 0.000 description 2
- 235000011149 sulphuric acid Nutrition 0.000 description 2
- 239000002699 waste material Substances 0.000 description 2
- 239000011667 zinc carbonate Substances 0.000 description 2
- 229910000010 zinc carbonate Inorganic materials 0.000 description 2
- NWONKYPBYAMBJT-UHFFFAOYSA-L zinc sulfate Chemical compound [Zn+2].[O-]S([O-])(=O)=O NWONKYPBYAMBJT-UHFFFAOYSA-L 0.000 description 2
- 229910000368 zinc sulfate Inorganic materials 0.000 description 2
- KEQXNNJHMWSZHK-UHFFFAOYSA-L 1,3,2,4$l^{2}-dioxathiaplumbetane 2,2-dioxide Chemical compound [Pb+2].[O-]S([O-])(=O)=O KEQXNNJHMWSZHK-UHFFFAOYSA-L 0.000 description 1
- 229910017626 NH4Fe(SO4)2 Inorganic materials 0.000 description 1
- 229910000503 Na-aluminosilicate Inorganic materials 0.000 description 1
- VEQPNABPJHWNSG-UHFFFAOYSA-N Nickel(2+) Chemical compound [Ni+2] VEQPNABPJHWNSG-UHFFFAOYSA-N 0.000 description 1
- BPQQTUXANYXVAA-UHFFFAOYSA-N Orthosilicate Chemical compound [O-][Si]([O-])([O-])[O-] BPQQTUXANYXVAA-UHFFFAOYSA-N 0.000 description 1
- 229910000831 Steel Inorganic materials 0.000 description 1
- 241001062472 Stokellia anisodon Species 0.000 description 1
- PNEYBMLMFCGWSK-UHFFFAOYSA-N aluminium oxide Inorganic materials [O-2].[O-2].[O-2].[Al+3].[Al+3] PNEYBMLMFCGWSK-UHFFFAOYSA-N 0.000 description 1
- XGGLLRJQCZROSE-UHFFFAOYSA-K ammonium iron(iii) sulfate Chemical compound [NH4+].[Fe+3].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O XGGLLRJQCZROSE-UHFFFAOYSA-K 0.000 description 1
- 229910052924 anglesite Inorganic materials 0.000 description 1
- 229910052925 anhydrite Inorganic materials 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 239000004566 building material Substances 0.000 description 1
- 238000001354 calcination Methods 0.000 description 1
- OSGAYBCDTDRGGQ-UHFFFAOYSA-L calcium sulfate Chemical compound [Ca+2].[O-]S([O-])(=O)=O OSGAYBCDTDRGGQ-UHFFFAOYSA-L 0.000 description 1
- 229910001429 cobalt ion Inorganic materials 0.000 description 1
- XLJKHNWPARRRJB-UHFFFAOYSA-N cobalt(2+) Chemical compound [Co+2] XLJKHNWPARRRJB-UHFFFAOYSA-N 0.000 description 1
- 229910052681 coesite Inorganic materials 0.000 description 1
- 239000002131 composite material Substances 0.000 description 1
- 239000012141 concentrate Substances 0.000 description 1
- 229910000365 copper sulfate Inorganic materials 0.000 description 1
- ARUVKPQLZAKDPS-UHFFFAOYSA-L copper(II) sulfate Chemical compound [Cu+2].[O-][S+2]([O-])([O-])[O-] ARUVKPQLZAKDPS-UHFFFAOYSA-L 0.000 description 1
- 229910052593 corundum Inorganic materials 0.000 description 1
- 229910052906 cristobalite Inorganic materials 0.000 description 1
- 238000000354 decomposition reaction Methods 0.000 description 1
- KZHJGOXRZJKJNY-UHFFFAOYSA-N dioxosilane;oxo(oxoalumanyloxy)alumane Chemical compound O=[Si]=O.O=[Si]=O.O=[Al]O[Al]=O.O=[Al]O[Al]=O.O=[Al]O[Al]=O KZHJGOXRZJKJNY-UHFFFAOYSA-N 0.000 description 1
- 238000003912 environmental pollution Methods 0.000 description 1
- 229910052840 fayalite Inorganic materials 0.000 description 1
- 229910001448 ferrous ion Inorganic materials 0.000 description 1
- 238000011049 filling Methods 0.000 description 1
- 238000010304 firing Methods 0.000 description 1
- PCHJSUWPFVWCPO-UHFFFAOYSA-N gold Chemical compound [Au] PCHJSUWPFVWCPO-UHFFFAOYSA-N 0.000 description 1
- 229910052737 gold Inorganic materials 0.000 description 1
- 239000010931 gold Substances 0.000 description 1
- XLYOFNOQVPJJNP-ZSJDYOACSA-N heavy water Substances [2H]O[2H] XLYOFNOQVPJJNP-ZSJDYOACSA-N 0.000 description 1
- XLYOFNOQVPJJNP-UHFFFAOYSA-M hydroxide Chemical compound [OH-] XLYOFNOQVPJJNP-UHFFFAOYSA-M 0.000 description 1
- DALUDRGQOYMVLD-UHFFFAOYSA-N iron manganese Chemical compound [Mn].[Fe] DALUDRGQOYMVLD-UHFFFAOYSA-N 0.000 description 1
- PNXOJQQRXBVKEX-UHFFFAOYSA-N iron vanadium Chemical compound [V].[Fe] PNXOJQQRXBVKEX-UHFFFAOYSA-N 0.000 description 1
- RUTXIHLAWFEWGM-UHFFFAOYSA-H iron(3+) sulfate Chemical compound [Fe+3].[Fe+3].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O RUTXIHLAWFEWGM-UHFFFAOYSA-H 0.000 description 1
- SZVJSHCCFOBDDC-UHFFFAOYSA-N iron(II,III) oxide Inorganic materials O=[Fe]O[Fe]O[Fe]=O SZVJSHCCFOBDDC-UHFFFAOYSA-N 0.000 description 1
- JEIPFZHSYJVQDO-UHFFFAOYSA-N iron(III) oxide Inorganic materials O=[Fe]O[Fe]=O JEIPFZHSYJVQDO-UHFFFAOYSA-N 0.000 description 1
- 229910000360 iron(III) sulfate Inorganic materials 0.000 description 1
- 229910052745 lead Inorganic materials 0.000 description 1
- JQJCSZOEVBFDKO-UHFFFAOYSA-N lead zinc Chemical compound [Zn].[Pb] JQJCSZOEVBFDKO-UHFFFAOYSA-N 0.000 description 1
- 239000011572 manganese Substances 0.000 description 1
- 229910044991 metal oxide Inorganic materials 0.000 description 1
- 150000004706 metal oxides Chemical class 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 229910052863 mullite Inorganic materials 0.000 description 1
- 229910052664 nepheline Inorganic materials 0.000 description 1
- 239000010434 nepheline Substances 0.000 description 1
- 229910001453 nickel ion Inorganic materials 0.000 description 1
- 239000010453 quartz Substances 0.000 description 1
- 229910052761 rare earth metal Inorganic materials 0.000 description 1
- 238000011160 research Methods 0.000 description 1
- 238000002791 soaking Methods 0.000 description 1
- 239000000429 sodium aluminium silicate Substances 0.000 description 1
- 235000012217 sodium aluminium silicate Nutrition 0.000 description 1
- URGAHOPLAPQHLN-UHFFFAOYSA-N sodium aluminosilicate Chemical compound [Na+].[Al+3].[O-][Si]([O-])=O.[O-][Si]([O-])=O URGAHOPLAPQHLN-UHFFFAOYSA-N 0.000 description 1
- 239000007790 solid phase Substances 0.000 description 1
- 239000002904 solvent Substances 0.000 description 1
- 239000010959 steel Substances 0.000 description 1
- 229910052682 stishovite Inorganic materials 0.000 description 1
- 239000000126 substance Substances 0.000 description 1
- 231100000331 toxic Toxicity 0.000 description 1
- 230000002588 toxic effect Effects 0.000 description 1
- 229910052905 tridymite Inorganic materials 0.000 description 1
- 229910052844 willemite Inorganic materials 0.000 description 1
- 229910001845 yogo sapphire Inorganic materials 0.000 description 1
- UGZADUVQMDAIAO-UHFFFAOYSA-L zinc hydroxide Chemical compound [OH-].[OH-].[Zn+2] UGZADUVQMDAIAO-UHFFFAOYSA-L 0.000 description 1
- 229940007718 zinc hydroxide Drugs 0.000 description 1
- 229910021511 zinc hydroxide Inorganic materials 0.000 description 1
- 229960001763 zinc sulfate Drugs 0.000 description 1
- 239000011686 zinc sulphate Substances 0.000 description 1
Images
Classifications
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/006—Wet processes
- C22B7/008—Wet processes by an alkaline or ammoniacal leaching
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01B—NON-METALLIC ELEMENTS; COMPOUNDS THEREOF; METALLOIDS OR COMPOUNDS THEREOF NOT COVERED BY SUBCLASS C01C
- C01B33/00—Silicon; Compounds thereof
- C01B33/113—Silicon oxides; Hydrates thereof
- C01B33/12—Silica; Hydrates thereof, e.g. lepidoic silicic acid
- C01B33/18—Preparation of finely divided silica neither in sol nor in gel form; After-treatment thereof
- C01B33/186—Preparation of finely divided silica neither in sol nor in gel form; After-treatment thereof from or via fluosilicic acid or salts thereof by a wet process
-
- 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
- C22B1/00—Preliminary treatment of ores or scrap
- C22B1/02—Roasting 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
- C22B13/00—Obtaining lead
- C22B13/04—Obtaining lead by wet processes
- C22B13/045—Recovery from waste materials
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B15/00—Obtaining copper
- C22B15/0063—Hydrometallurgy
- C22B15/0065—Leaching or slurrying
-
- 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
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/20—Obtaining zinc otherwise than by distilling
- C22B19/24—Obtaining zinc otherwise than by distilling with leaching with alkaline solutions, e.g. ammonia
-
- 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
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/30—Obtaining zinc or zinc oxide from metallic residues or scraps
-
- 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
Abstract
The invention belongs to the technical field of solid waste recycling, and particularly relates to a method for reducing and utilizing multi-metal tailings. The method comprises the steps of sodium carbonate proportioning roasting, alkali liquor leaching, ammonium bisulfate blending roasting, water leaching, ammoniojarosite precipitation, heavy metal ion precipitation, aluminum extraction, silicon extraction to prepare white carbon black or silica gel, recovery and zinc extraction from the heavy metal ion precipitation, zinc salt hydrothermal reaction to prepare zinc oxide and the like. The method for reducing and utilizing the multi-metal tailings not only extracts aluminum and silicon from the multi-metal tailings to prepare the photocatalyst and the ammonioiarosite, but also fixes heavy metal ions such as lead, copper, Cd and the like with high toxicity in the tailings, and acid and alkali used in the whole process can be recycled in the process. The method can realize high value-added utilization of the multi-metal tailings and realize harmless and reduction treatment of the multi-metal tailings.
Description
Technical Field
The invention belongs to the technical field of solid waste recycling, and particularly relates to a method for reducing and utilizing multi-metal tailings.
Background
The tailings refer to solid wastes left after ores are crushed and ground to smelt 'useful component' concentrates under specific economic and technical conditions. The mining and dressing recovery rate of the Chinese mining industry to the target minerals is 20 percent lower than that of developed countries in the west. Wherein the average recovery rate of the ferrous manganese iron metal mine is only 65%, and the comprehensive recovery rate of the nonferrous metal mine is only 60% -70%; the comprehensive recovery utilization rate of lead and zinc ores in a dressing plant is less than 30 percent. In addition, 80% of Chinese mineral resources are associated minerals, and particularly, although main mineral elements in the minerals are recovered to a certain extent, associated valuable elements are often recovered little or even not, so that a large amount of valuable resources except the main mineral elements are remained in tailings and are wasted. Only by taking ladle steel as an example, the proportion of rare earth elements contained in 1.8 million tons of currently discharged tailing slag is higher than that of Baiyunebo raw ore except bailey. China is a large country for mining mineral resources, and the amount of tailings discharged every year is very large. According to Chinese mineral resource saving and comprehensive utilization reports (2015), the annual discharge amount of tailings is up to more than 15 hundred million tons only in 5 years from 2011 to 2015, and at present, 146 hundred million tons of tailings are accumulated and stockpiled in a tailing dam in China, and most of tailings are accumulated in the tailing dam. According to the national mineral resource saving and comprehensive utilization report (2016), the national tailing stock reaches 146 hundred million tons by 2015. According to the current tailing discharge increment, the stock of the tailings in China reaches 200 hundred million tons by the end of 2018. The large amount of discharged and accumulated tailings not only occupy land, waste precious natural resources and seriously pollute the environment, but also have great potential safety hazard, and once a dam break accident occurs, huge environmental disasters, economic losses and casualties can be brought. How to solve the environmental and safety problems caused by the discharge of tailings and improve the comprehensive utilization rate of tailings is always a hot problem which is very concerned and urgently to be solved by relevant departments of relevant industries and countries.
At present, tailings are mainly used for secondary mineral separation, filling and mining, preparation of various building materials, agriculture, recovery and treatment of the environment and the like, but the utilization rate is low, particularly in China, the utilization rate only reaches 18.9% of the total amount of the tailings, and the economic value of the utilization is low. The reason is that the comprehensive utilization technology of the tailings capable of producing high value-added products is lacked. Therefore, the method solves the problem of large-scale stockpiling of the tailings, prolongs the industrial chain of metal tailing products, improves the resource utilization rate of the metal tailings, urgently needs to develop a new comprehensive utilization process technology of the metal tailings, and realizes substantial breakthrough and upgrade from metal tailing raw materials to new resource utilization products.
In the comprehensive utilization method of multi-metal tailings (including but not limited to gold, silver, lead and zinc ore tailings, lead, zinc and copper ore tailings, copper and silver ore tailings, silver, lead and zinc ore tailings, copper multi-metal ore tailings and the like), the prior art usually only recycles one or a few metal elements in the tailings, and the cost is higher and the economic benefit is not obvious. In particular to multi-metal tailings, a primary deposit of the multi-metal tailings often contains three or more ore-forming metal elements, other associated metal elements are more in types and higher in content, and more metal elements with higher grade are still added in the tailings after ore dressing. If the recovery, enrichment and utilization of the residual valuable metal elements in the multi-metal tailings cannot be realized, the heavy metals in the tailings pollute the environment and cause a great deal of resource waste and serious economic loss.
Disclosure of Invention
Technical problem to be solved
In order to solve the problems in the prior art, the invention provides a method for reducing and utilizing multi-metal tailings, which can recover and separate silicon and aluminum in various metal tailings, separate and enrich metal elements, and realize reduction and recycling of the metal tailings.
(II) technical scheme
In order to achieve the purpose, the invention adopts the main technical scheme that:
in one aspect, the invention provides a method for reducing and utilizing multi-metal tailings, which comprises the following steps:
step 1: grinding the metal tailings into 150-mesh and 250-mesh mineral powder, uniformly mixing the mineral powder and sodium carbonate, and roasting at the roasting temperature of 800-mesh and 900 ℃ for 0.5-3h to obtain roasted clinker;
step 2: dissolving out the roasted clinker by using an aqueous solution of sodium hydroxide, and then carrying out solid-liquid separation, wherein a liquid phase contains sodium silicate and sodium metaaluminate to obtain pre-desiliconized solid slag;
and step 3: grinding the solid slag into 150-mesh powder with the particle size of 250 meshes, mixing the powder with ammonium bisulfate, roasting to obtain a roasted product, and absorbing tail gas generated in the roasting process by dilute sulfuric acid or water to prepare ammonium sulfate or ammonium bisulfate;
and 4, step 4: adding water to the roasted product obtained in the step 3 for leaching, and carrying out solid-liquid separation under the heating condition to obtain leaching slag and leaching liquid, wherein the leaching slag contains lead sulfate and silicon slag;
and 5: aerating the leachate at a rate of 0.8-1.2L/L per minute for 20-40min to remove Fe2+Is oxidized into Fe3+(ii) a Adding an ammonium bicarbonate solution into the leachate to adjust the pH value of the leachate, controlling the temperature of the solution to be 90-95 ℃, firstly controlling the pH value of the solution to be 1.5-2 to carry out iron precipitation reaction, and filtering after the iron precipitation is finished to obtain the iron-containing composite materialPrecipitating the filtrate and ammonium jarosite;
step 6: adjusting the pH of the filtrate to precipitate metal ions: continuously and slowly adding a sodium hydroxide solution into the filtrate obtained in the step (5), adjusting the pH value of the filtrate to 9.5-10.5, controlling the temperature to be 20-35 ℃, and filtering to obtain filter residue and filtrate; wherein the filter residue mainly contains Cu (OH)2And Zn (OH)2The filtrate contains metaaluminate ions;
and 7: combining the liquid phase obtained by separation in the step 2 with the filtrate obtained in the step 6, and adjusting the pH value to be between 5.5 and 8 by using weak acid salt to obtain an aluminum hydroxide precipitate, wherein the liquid phase contains a sodium silicate solution;
and 8: dissolving the filter residue obtained in the step 6 in a caustic soda solution, adjusting the pH value to be more than or equal to 12, allowing zincate ions to enter a liquid phase, and filtering for solid-liquid separation to obtain filtrate and filter residue; the filter residue contains Cu (OH)2Fixing the copper ions for further processing; adding the filtrate into a hydrothermal reaction kettle, carrying out hydrothermal reaction for 5-12 h at the temperature of 150-220 ℃, and separating to obtain zinc oxide precipitate;
or adding dilute sulfuric acid and zinc powder into the filter residue obtained in the step 6, reacting and filtering, adjusting the pH of the filtrate to be more than or equal to 11, transferring the filtrate into a hydrothermal reaction kettle, carrying out hydrothermal reaction for 5-12 hours at the temperature of 150-220 ℃, and separating to obtain zinc oxide precipitate; the filter residue contains copper.
According to a preferred embodiment of the invention, in the step 1, the mass ratio of the mineral powder to the sodium carbonate is 1: 0.8-2.5.
Wherein sodium carbonate can react with aluminum and silicon in tailings to generate nepheline, and CO is released in the reaction process2The iron tailing active clinker has good solubility and a loose structure, more than 90% of aluminum and silicon in the tailings can be effectively activated, and an alkaline leaching solution environment is also provided in the step 2; the sodium carbonate can also promote the decomposition of fayalite, mullite and quartz. The research results of thermodynamics and pure substance tests show that sodium carbonate can react with silicon and aluminum to preferentially generate sodium aluminosilicate.
According to a preferred embodiment of the invention, in the step 2, the concentration of the sodium hydroxide solution used for dissolution by using the sodium hydroxide solution is 0.05-0.3 mol/L, the solid-liquid mass ratio during dissolution is 1: 3-5, the dissolution temperature is 80-95 ℃, and the dissolution reaction time is 0.5-3 h.
According to the preferred embodiment of the invention, in step 3, the amount of ammonium bisulfate is 1.5-2 times of the theoretical amount required for completely reacting zinc, iron, copper and aluminum in the powder, the powder is roasted for 1-4h at the temperature of 350-. The dilute sulfuric acid can be used for preparing reaction raw material ammonium bisulfate.
According to the preferred embodiment of the invention, in the step 4, the water is added according to the solid-to-liquid ratio of 1:5-6, stirred and dissolved for 20-60min at the temperature of 50-90 ℃, and solid-liquid separation is carried out to obtain leaching slag and leaching liquid, wherein the leaching slag mainly contains lead sulfate and silicon slag.
According to the preferred embodiment of the invention, the leached slag containing lead sulfate and silicon slag obtained in the step 4 is dissolved in caustic soda solution, and lead sulfate and water glass are obtained by separation; adding dilute sulfuric acid or hydrochloric acid into water glass, standing to obtain hydrated silicic acid gel, solidifying, washing with water to remove electrolyte ions dissolved therein, drying to obtain silica gel, and converting the water solution into sodium sulfate or sodium chloride solution.
According to a preferred embodiment of the present invention, the ozone in step 5 can be replaced by hydrogen peroxide.
According to the preferred embodiment of the invention, dilute sulphuric acid or hydrochloric acid is added into the sodium silicate solution in the step 7 and is kept stand to obtain hydrous silicic acid gel which is solidified, electrolyte ions dissolved in the hydrous silicic acid gel are removed by water washing, and silica gel is obtained after drying, wherein the washing liquid is a solution of sodium sulfate or sodium chloride; or: introducing CO into the water glass solution2And (3) fully carbonizing the water glass to generate micro-nano white carbon black, simultaneously adding the active lime emulsion into the obtained sodium carbonate solution to perform causticization reaction, and then filtering and separating to obtain calcium carbonate solid and a sodium hydroxide solution, wherein the sodium hydroxide solution is recycled in the step 2.
According to a preferred embodiment of the present invention, the weak acid salt used in step 7 is ammonium sulfate or ammonium bisulfate.
According to a preferred embodiment of the invention, in the step 8, in the hydrothermal reaction, 8-10% by mass of PEG20000 or SDBS solution is added, the amount of the PEG20000 or SDBS solution is 4.0-8.0% of the total volume of the reaction system in the hydrothermal reaction kettle, the mixture is uniformly stirred and then reacts at 180-210 ℃ for 6-10h to generate zinc oxide precipitate, and the zinc oxide precipitate is dried to obtain petaloid zinc oxide powder serving as a photocatalyst; a large amount of sulfate in the filtrate after the hydrothermal reaction can be used for preparing ammonium sulfate or ammonium bisulfate and sodium sulfate, and can be recycled in the whole process flow.
In the method for reducing and utilizing the polymetallic tailings, the tailings to be treated include, but are not limited to, one or more of tailings mainly containing iron, copper, zinc, lead, aluminum, silicon and the like, such as lead-zinc tailings, copper-containing tailings, bauxite tailings, magnetite tailings and the like. The process line designed by the invention is mainly used for obtaining the aluminum, the silicon, the iron and the zinc in the tailings, fixing toxic heavy metals such as copper, lead, cadmium and the like contained in the tailings, realizing the reduction of the tailings and preventing a large amount of tailings from accumulating to irreversibly damage farmlands and water quality.
(III) advantageous effects
The invention has the beneficial effects that:
the method for reducing and utilizing the polymetallic tailings not only prepares silica gel or white carbon black from the polymetallic tailings and extracts aluminum to prepare photocatalyst zinc oxide and jarosite precipitates, but also enriches and fixes heavy metal ions such as lead, copper, Cd and the like with high toxicity in the tailings, and acid and alkali used in the whole process can be recycled in the process. The method has strong applicability, controllable process and good economic benefit, can realize high value-added utilization of the multi-metal tailings, reduce the pollution of heavy metals in the tailings on soil and water around a mining area, and simultaneously reduce the tailings.
Drawings
FIG. 1 is a schematic flow chart of a preferred embodiment of the present invention; the dotted line in the figure indicates the solid phase precipitation and the solid line is the filtrate or liquid phase.
Detailed Description
For the purpose of better explaining the present invention and to facilitate understanding, the present invention will be described in detail by way of specific embodiments with reference to the accompanying drawings.
As shown in fig. 1, a method for reducing and utilizing multi-metal tailings according to a preferred embodiment of the present invention includes:
step 1: grinding the metal tailings into 150-mesh and 250-mesh mineral powder, uniformly mixing the mineral powder and sodium carbonate, and roasting at the roasting temperature of 800-mesh and 900 ℃ for 0.5-3h to obtain the roasted clinker.
Preferably, the mass ratio of the mineral powder to the sodium carbonate is 1: 0.8-2.5.
By mixing and roasting the mineral powder and the sodium carbonate, the obtained multi-metal tailing active clinker has good solubility, the roasted clinker becomes porous and has a loose structure, and more than 90% of Cu, Pb, Zn and Si in the clinker can be effectively activated.
Step 2: dissolving out the roasted clinker by using an aqueous solution of sodium hydroxide, and then carrying out solid-liquid separation, wherein the liquid phase contains sodium silicate and sodium metaaluminate to obtain the pre-desiliconized solid slag.
And (3) leaching by using sodium hydroxide, wherein part of aluminum and part of silicon are converted into metaaluminate and silicate, and enter a liquid phase.
Preferably, the concentration of the sodium hydroxide solution used for dissolution by using the sodium hydroxide solution is 0.05-0.3 mol/L, the solid-liquid mass ratio during dissolution is 1: 3-5, the dissolution temperature is 80-95 ℃, and the dissolution reaction time is 0.5-3 h.
And step 3: grinding the solid slag into 150-fold 250-mesh powder, mixing the powder with ammonium bisulfate, wherein the dosage of the ammonium bisulfate is 1.5-2 times of the theoretical amount required for completely reacting zinc, iron, copper and aluminum in the powder, roasting for 1-4h at the temperature of 350-fold 400 ℃, and absorbing tail gas generated in the roasting process by dilute sulfuric acid or water to prepare ammonium sulfate or ammonium bisulfate.
The chemical reactions that occur during calcination are as follows:
ZnCO3＝ZnO+CO2↑
ZnCO3+2NH4HSO4＝(NH4)2Zn(SO4)2+H2O↑+CO2↑
ZnO+2NH4HSO4＝(NH4)2Zn(SO4)2+H2O↑
Zn2SiO4+4NH4HSO4＝2(NH4)2Zn(SO4)2+2H2O↑+SiO2↑
Fe2O3+6NH4HSO4＝2(NH4)3Fe(SO4)3+3H2O↑
Al2O3+6NH4HSO4＝2(NH4)3Al(SO4)3+3H2O↑
(NH4)3Fe(SO4)3＝NH4Fe(SO4)2+2NH3↑+H2O↑+SO3↑
(NH4)2Zn(SO4)2＝ZnSO4+2NH3↑+H2O↑+SO3↑
NH4HSO4＝H2SO4+NH3↑
CaCO3+H2SO4＝CaSO4+H2O↑+CO2↑
PbO+H2SO4＝PbSO4+H2O↑
during the firing process, a large amount of gas is generated, making the fired mass porous and loose. The reaction temperature of roasting zinc oxide ore by ammonium bisulfate is 350-400 ℃, so that the extraction rate of zinc can be ensured. After the ammonium bisulfate is decomposed, sulfuric acid can be generated, and under the conditions of roasting and high temperature, the generated sulfuric acid has strong oxidability, can destroy acid radicals or ion groups in minerals, has strong acidity after being dissolved by adding water, and can be used for acid leaching of metal elements in metal oxides.
Wherein, preferably, the dosage of the ammonium bisulfate is 1.5 times of the theoretical amount required for completely reacting the zinc, the iron, the copper and the aluminum in the powder, the roasting is carried out for 2 hours at the temperature of 350 ℃, and tail gas generated in the roasting process is absorbed by dilute sulfuric acid or water to prepare the ammonium sulfate or the ammonium bisulfate. The dilute sulfuric acid can be used for preparing reaction raw material ammonium bisulfate and can be recycled in the process flow.
And 4, step 4: and (3) adding water to leach the roasted product obtained in the step (3) with a solid-to-liquid ratio of 1:5-6, stirring and dissolving at 50-90 ℃ for 20-60min, and carrying out solid-liquid separation to obtain leaching residues and leaching liquid, wherein the leaching residues contain lead sulfate and silicon slag.
Preferably, water is added for leaching according to the solid-to-liquid ratio of 1:5, stirring and dissolving are carried out for 30min at the temperature of 60 ℃, solid-liquid separation is carried out, and leaching residue and leaching liquid are obtained, wherein the leaching residue mainly contains lead sulfate and silicon slag.
The leachate mainly contains sulfate, including ferric sulfate, copper sulfate, zinc sulfate and the like, and the slag also contains partial silica slag besides insoluble lead sulfate.
And the leached slag containing lead sulfate and silicon slag can be further dissolved in caustic soda solution, and lead sulfate and water glass are obtained through separation. The water glass can be converted into white carbon black or silica gel. If dilute sulphuric acid or hydrochloric acid is added into water glass and is kept stand to obtain hydrous silicic acid gel for solid state, electrolyte ions dissolved in the hydrous silicic acid gel are removed by water washing, silica gel is obtained after drying, and the washing liquid is sodium sulfate or sodium chloride solution.
And 5: aerating the leachate at a rate of 0.8-1.2L/L per minute for 20-40min to remove Fe2+Is oxidized into Fe3+(ii) a Adding an ammonium bicarbonate solution into the leachate to adjust the pH value of the leachate, controlling the temperature of the solution to be 90-95 ℃, firstly controlling the pH value of the solution to be 1.5-2 to carry out iron precipitation reaction, and after the iron precipitation is finished, filtering to obtain a filtrate and an ammonium jarosite precipitate.
The purpose of the ozone introduction is mainly to convert various sub-metal ions into higher valence states, such as Fe2+Is oxidized into Fe3+. However, in the case of containing copper ions or cobalt ions, Fe is hardly present2+However, in order to convert all ferrous ions into ferric ions, a certain amount of ozone can be introduced or hydrogen peroxide can be added to facilitate the subsequent precipitation of the vanadium iron.
Step 6: continuously and slowly adding sodium hydroxide solution into the filtrate obtained in the step 5, adjusting the pH value of the filtrate to 9.5-10.5, and controllingThe preparation temperature is 20-35 ℃, and filtration is carried out to obtain filter residue and filtrate; wherein the filter residue mainly contains Cu (OH)2And Zn (OH)2The filtrate contains metaaluminate ion and a small amount of sodium silicate. Wherein most of nickel ions, Cd ions, Mn ions, etc. are precipitated.
In adjusting the pH by adding sodium hydroxide solution, reference may be made to the respective metal precipitates and solvent pH ranges as in table 1 below.
Table 1: precipitation pH of different metal ions
And 7: and (3) combining the liquid phase obtained by separation in the step (2) and the filtrate obtained in the step (6), and adjusting the pH value to be between 5.5 and 8 by using weak acid salt to obtain an aluminum hydroxide precipitate, wherein the liquid phase contains a sodium silicate solution.
Adding dilute sulfuric acid or hydrochloric acid into a sodium silicate solution, standing to obtain hydrous silicic acid gel for solid state, washing with water to remove electrolyte ions dissolved in the hydrous silicic acid gel, and drying to obtain silica gel, wherein a washing solution is a solution of sodium sulfate or sodium chloride; or: introducing CO into the water glass solution2And (3) fully carbonizing the water glass to generate micro-nano white carbon black, simultaneously adding the active lime emulsion into the obtained sodium carbonate solution to perform causticization reaction, and then filtering and separating to obtain calcium carbonate solid and a sodium hydroxide solution, wherein the sodium hydroxide solution is recycled in the step 2. Wherein the weak acid salt for adjusting pH is ammonium bisulfate or ammonium sulfate.
And 8: dissolving the filter residue obtained in the step 6 in a caustic soda solution with the pH value of more than 12, allowing zincate ions to enter a liquid phase, and performing solid-liquid separation; and adding the filtrate into a hydrothermal reaction kettle, carrying out hydrothermal reaction for 5-12 h at the temperature of 150-220 ℃, and separating to obtain zinc oxide precipitate. The filter residue after solid-liquid separation contains Cu (OH)2The copper ions are fixed for further processing.
Wherein, in the hydrothermal reaction process, PEG20000 or SDBS solution with the mass percentage concentration of 8-10% can be added, the dosage of the PEG20000 or SDBS solution is 4.0-8.0% of the total volume of the reaction system in the hydrothermal reaction kettle, the mixture is uniformly stirred and then reacts for 6-10h at 180-210 ℃ to generate zinc oxide precipitate, and the zinc oxide precipitate is dried to obtain zinc oxide powder with petal-shaped structure which can be used as photocatalyst; a large amount of sulfate in the filtrate after the hydrothermal reaction can be used for preparing ammonium sulfate or ammonium bisulfate and sodium sulfate, and can be recycled in the whole process flow.
The following are specific examples of the present invention.
Example 1
The embodiment is used for reducing and treating bauxite tailings, and comprises the following steps:
step 1: grinding bauxite tailings into mineral powder of 200 meshes, uniformly mixing the mineral powder and sodium carbonate, and roasting at 900 ℃ for 2 hours to obtain roasted clinker; the mass ratio of the mineral powder to the sodium carbonate is 1: 2.
Step 2: dissolving out the roasted clinker by using 0.3mol/L aqueous solution of sodium hydroxide, wherein the solid-liquid mass ratio is 1:4, the dissolving-out temperature is 95 ℃, the dissolving-out time is 2 hours, and carrying out solid-liquid separation, wherein the liquid phase contains sodium silicate and sodium metaaluminate to obtain the pre-desiliconized solid slag.
And step 3: grinding the solid slag into powder of 200 meshes, mixing the powder with ammonium bisulfate, wherein the dosage of the ammonium bisulfate is 1.5 times of the theoretical quantity required for completely reacting zinc, iron, copper and aluminum in the powder, roasting for 2 hours at 350 ℃, and absorbing tail gas generated in the roasting process by dilute sulfuric acid or water to prepare ammonium sulfate or ammonium bisulfate.
And 4, step 4: and (3) adding water to leach the roasted product obtained in the step (3) with a solid-to-liquid ratio of 1:5, stirring and dissolving at 60 ℃ for 30min, and carrying out solid-liquid separation to obtain leaching slag and leaching liquid, wherein the leaching slag contains lead sulfate and silicon slag.
And the leached slag containing lead sulfate and silicon slag can be further dissolved in caustic soda solution, and lead sulfate and water glass are obtained through separation.
And 5: aerating the leachate with ozone at normal pressure of 1L/min for 30min to remove Fe2+Is oxidized into Fe3+(ii) a Adding an ammonium bicarbonate solution into the leachate to adjust the pH value of the leachate, controlling the temperature of the solution to be 90 ℃, firstly controlling the pH value of the solution to be 1.5-2 to carry out iron precipitation reaction, after the iron precipitation is finished,filtering to obtain filtrate and precipitate of ammonioiarosite.
Step 6: continuously and slowly adding a sodium hydroxide solution into the filtrate obtained in the step (5), adjusting the pH value of the filtrate to 9.5-10.5, controlling the temperature to be 35 ℃, and filtering to obtain filter residues and filtrate; wherein the filter residue mainly contains Cu (OH)2And Zn (OH)2The filtrate contains metaaluminate ion and a small amount of sodium silicate.
And 7: combining the liquid phase separated in step 2 with the filtrate of step 6, adding ammonium bisulfate, and adjusting pH to 5.5-8 to obtain aluminum hydroxide precipitate and liquid phase containing sodium silicate solution.
Continuously introducing CO into the sodium silicate solution2And (3) fully carbonizing the water glass to generate micro-nano white carbon black, adding the active lime emulsion into the obtained sodium carbonate solution to perform causticization reaction, and then filtering and separating to obtain calcium carbonate solid and a sodium hydroxide solution, wherein the sodium hydroxide solution can be reused in the step 2 after concentration adjustment.
And 8: and (3) dissolving the filter residue obtained in the step (6) in a caustic soda solution with the pH value of more than 12, wherein most of zinc in the zinc hydroxide enters a liquid phase by zincate ions, and carrying out solid-liquid separation. The filter residue is mainly Cu (OH)2And hydroxide precipitates of other non-amphoteric metals. Adding the filtrate into a hydrothermal reaction kettle, adding 8 mass percent of PEG20000 solution, wherein the dosage of the PEG20000 solution is 5% of the total volume of a reaction system in the hydrothermal reaction kettle, uniformly stirring, and reacting at 180 ℃ for 6h to generate the zinc oxide powder with the petal-shaped structure, and the zinc oxide powder has a large specific surface area and can be used as a photocatalyst.
A large amount of sulfate in the filtrate after the hydrothermal reaction can be used for preparing ammonium sulfate or ammonium bisulfate and sodium sulfate and can be recycled in required links in the whole process flow.
By testing the method in a laboratory scale, the bauxite tailings can be reduced by more than 65 percent, aluminum and silicon elements in the bauxite tailings are recovered, white carbon black or silica gel is prepared, ammonium jarosite is prepared, and lead, copper, cadmium and the like which have great effects on environmental pollution are fixed.
Example 2
This example was used for the germany copper tailings, comprising the following steps:
step 1: grinding the copper tailings into mineral powder of 200 meshes, uniformly mixing the mineral powder and sodium carbonate, and roasting at 850 ℃ for 2.5 hours to obtain roasted clinker; the mass ratio of the mineral powder to the sodium carbonate is 1: 1.5.
Step 2: dissolving out the roasted clinker by using 0.25mol/L aqueous solution of sodium hydroxide, wherein the solid-liquid mass ratio is 1:6, the dissolving-out temperature is 90 ℃, the dissolving-out time is 2 hours, and carrying out solid-liquid separation, wherein the liquid phase contains sodium silicate and sodium metaaluminate to obtain the pre-desiliconized solid slag.
And step 3: grinding the solid slag into powder of 200 meshes, mixing the powder with ammonium bisulfate, wherein the dosage of the ammonium bisulfate is 2 times of the theoretical quantity required for completely reacting zinc, iron, copper and aluminum in the powder, roasting for 3 hours at 400 ℃, and absorbing tail gas generated in the roasting process by dilute sulfuric acid or water to prepare ammonium sulfate or ammonium bisulfate.
And 4, step 4: and (3) adding water to leach the roasted product obtained in the step (3) with a solid-to-liquid ratio of 1:6, stirring and dissolving at 90 ℃ for 60min, and carrying out solid-liquid separation to obtain leaching slag and leaching liquid, wherein the leaching slag contains lead sulfate and silicon slag. And the leached slag containing lead sulfate and silicon slag can be further dissolved in caustic soda solution, and lead sulfate and water glass are obtained through separation.
And 5: aerating the leachate with ozone at a rate of 0.8L/L per minute under normal pressure for 20min to remove Fe2+Is oxidized into Fe3+(ii) a Adding an ammonium bicarbonate solution into the leachate to adjust the pH value of the leachate, controlling the temperature of the solution to be 90 ℃, firstly controlling the pH value of the solution to be 1.5-2 to carry out iron precipitation reaction, and after the iron precipitation is finished, filtering to obtain a filtrate and an ammonium jarosite precipitate.
Step 6: continuously and slowly adding a sodium hydroxide solution into the filtrate obtained in the step (5), adjusting the pH of the filtrate to 9.5-10.5, controlling the temperature to be 30 ℃, and filtering to obtain filter residue and filtrate; wherein the filter residue mainly contains Cu (OH)2And Zn (OH)2The filtrate contains metaaluminate ion and a small amount of sodium silicate.
And 7: and (3) combining the liquid phase obtained by separation in the step (2) and the filtrate obtained in the step (6), and adjusting the pH value to be between 5.5 and 8 by using ammonium sulfate to obtain an aluminum hydroxide precipitate, wherein the liquid phase contains a sodium silicate solution.
And 8: and (4) adding dilute sulfuric acid into the filter residue obtained in the step (6) and soaking metal ions into acid liquor. Measuring the concentrations of copper ions, Cr ions and Cd ions in the acid liquor, estimating the total molar weight, adding 1.05-1 times of zinc powder to replace the oxidizing metals such as copper and Cd, filtering and recovering the simple metal. Adding a caustic soda solution into the filtrate, transferring the filtrate into a hydrothermal reaction kettle, carrying out hydrothermal reaction for 5-12 h at the temperature of 150-220 ℃, and separating to obtain zinc oxide precipitate. A large amount of sulfate in the filtrate after the hydrothermal reaction can be used for preparing ammonium sulfate or ammonium bisulfate and sodium sulfate, and can be recycled in the whole process flow.
The inventor tests the method on a laboratory scale, and tests show that the method can reduce the copper tailings by 70 percent, recover aluminum, silicon and iron elements in the copper tailings, prepare white carbon black or silica gel, and fix heavy metals of lead, copper, cadmium and the like to avoid the pollution of the heavy metals on the soil and water quality around a mineral area.
Finally, it should be noted that: the above-mentioned embodiments are only used for illustrating the technical solution of the present invention, and not for limiting the same; although the present invention has been described in detail with reference to the foregoing embodiments, it will be understood by those of ordinary skill in the art that: it is to be understood that modifications may be made to the above-described embodiments, or equivalents may be substituted for some or all of their features, without departing from the scope of the embodiments of the present invention.
Claims (9)
1. A method for reducing and utilizing multi-metal tailings is characterized by comprising the following steps:
step 1: sodium carbonate blending roasting: grinding the metal tailings into 150-mesh and 250-mesh mineral powder, uniformly mixing the mineral powder and sodium carbonate, and roasting at the roasting temperature of 800-mesh and 900 ℃ for 0.5-3h to obtain roasted clinker;
step 2: alkaline leaching: dissolving out the roasted clinker by using an aqueous solution of sodium hydroxide, and then carrying out solid-liquid separation, wherein a liquid phase contains sodium silicate and sodium metaaluminate to obtain pre-desiliconized solid slag;
and step 3: and (3) blending and roasting ammonium bisulfate: grinding the solid slag into 150-mesh powder with the particle size of 250 meshes, mixing the powder with ammonium bisulfate, roasting to obtain a roasted product, and absorbing tail gas generated in the roasting process by dilute sulfuric acid or water to prepare ammonium sulfate or ammonium bisulfate;
and 4, step 4: adding water for leaching: adding water to the roasted product obtained in the step 3 for leaching, and carrying out solid-liquid separation under the heating condition to obtain leaching slag and leaching liquid, wherein the leaching slag contains lead sulfate and silicon slag;
and 5: oxidizing and depositing iron in the leaching solution: aerating the leachate at a rate of 0.8-1.2L/L per minute for 20-40min to remove Fe2+Is oxidized into Fe3+(ii) a Adding an ammonium bicarbonate solution into the leachate to adjust the pH value of the leachate, controlling the temperature of the solution to be 90-95 ℃, firstly controlling the pH value of the solution to be 1.5-2 to carry out iron precipitation reaction, and after the iron precipitation is finished, filtering to obtain a filtrate and an ammonium jarosite precipitate;
step 6: copper deposition: adjusting the pH of the filtrate to precipitate metal ions: continuously and slowly adding a sodium hydroxide solution into the filtrate obtained in the step (5), adjusting the pH value of the filtrate to 9.5-10.5, controlling the temperature to be 20-35 ℃, and filtering to obtain filter residue and filtrate; wherein the filter residue mainly contains Cu (OH)2And Zn (OH)2The filtrate contains metaaluminate ions;
and 7: aluminum deposition: combining the liquid phase obtained by separation in the step 2 with the filtrate obtained in the step 6, and adjusting the pH value to be between 5.5 and 8 by using weak acid salt to obtain an aluminum hydroxide precipitate, wherein the liquid phase contains a sodium silicate solution;
and 8: zinc extraction: dissolving the filter residue obtained in the step 6 in a caustic soda solution, adjusting the pH value to be more than or equal to 12, allowing zincate ions to enter a liquid phase, and filtering for solid-liquid separation to obtain filtrate and filter residue; the filter residue contains Cu (OH)2Fixing the copper ions for further processing; adding the filtrate into a hydrothermal reaction kettle, carrying out hydrothermal reaction for 5-12 h at the temperature of 150-220 ℃, and separating to obtain zinc oxide precipitate;
or adding dilute sulfuric acid and zinc powder into the filter residue obtained in the step 6, reacting and filtering, adjusting the pH of the filtrate to be more than or equal to 11, transferring the filtrate into a hydrothermal reaction kettle, carrying out hydrothermal reaction for 5-12 hours at the temperature of 150-220 ℃, and separating to obtain zinc oxide precipitate; the filter residue contains copper.
2. The method according to claim 1, wherein in the step 1, the mass ratio of the mineral powder to the sodium carbonate is 1: 0.8-2.5.
3. The method according to claim 1, wherein in the step 2, the concentration of the sodium hydroxide solution used for elution with the sodium hydroxide solution is 0.05 to 0.3mol/L, the solid-liquid mass ratio during elution is 1:3 to 5, the elution temperature is 80 to 95 ℃, and the elution reaction time is 0.5 to 3 hours.
4. The method as claimed in claim 1, wherein in step 3, the amount of ammonium bisulfate used is 1.5-2 times of the theoretical amount required for completely reacting zinc, iron, copper and aluminum in the powder, the powder is roasted for 1-4h at the temperature of 350-400 ℃, and tail gas generated in the roasting process is absorbed by dilute sulfuric acid or water to prepare ammonium sulfate or ammonium bisulfate.
5. The method according to claim 1, wherein in the step 4, the amount of water is 1:5-6 in a solid-to-liquid ratio, the leaching is carried out for 20-60min under stirring at 50-90 ℃, and solid-liquid separation is carried out to obtain leaching residue and leaching liquid, wherein the leaching residue mainly contains lead sulfate and silicon slag.
6. The method according to claim 1, characterized in that the leached slag containing lead sulfate and silicon slag obtained in the step 4 is dissolved in caustic soda solution and separated to obtain lead sulfate and water glass; adding dilute sulfuric acid or hydrochloric acid into water glass, standing to obtain hydrated silicic acid gel, solidifying, washing with water to remove electrolyte ions dissolved therein, drying to obtain silica gel, and converting the water solution into sodium sulfate or sodium chloride solution.
7. The method of claim 1, wherein the ozone in step 5 is replaced with hydrogen peroxide.
8. The method according to claim 1, wherein dilute sulfuric acid or hydrochloric acid is added to the sodium silicate solution of step 7 and left to stand to obtain a hydrous silicic acid gel to be solidified, electrolyte ions dissolved therein are removed by washing with water, and silica gel is obtained after drying, and the washing solution is a solution of sodium sulfate or sodium chloride;
or: introducing CO into the water glass solution2And (3) fully carbonizing the water glass to generate micro-nano white carbon black, simultaneously adding the active lime emulsion into the obtained sodium carbonate solution to perform causticization reaction, and then filtering and separating to obtain calcium carbonate solid and a sodium hydroxide solution, wherein the sodium hydroxide solution is recycled in the step 2.
9. The method according to claim 1, wherein in the step 8, a PEG20000 or SDBS solution with a mass percentage concentration of 8-10% is added during the hydrothermal reaction, the use amount of the PEG20000 or SDBS solution is 4.0% -8.0% of the total volume of the reaction system in the hydrothermal reaction kettle, the mixture is uniformly stirred and then reacts at 180-210 ℃ for 6-10h to generate zinc oxide precipitate, and the zinc oxide precipitate is dried to obtain the zinc oxide powder with the petal-shaped structure, which is used as the photocatalyst.
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| CN113122735A (en) * | 2021-04-02 | 2021-07-16 | 云南云铜锌业股份有限公司 | Alkali-process zinc powder combined smelting method |
| CN113830776A (en) * | 2021-10-18 | 2021-12-24 | 北京润捷浩达科技有限公司 | Method for recovering polymetallic crystal co-production water glass from copper-nickel sulfide ore tailings |
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