EP3172348A1 - Recovery of zinc and manganese from pyrometallurgy sludge or residues - Google Patents
Recovery of zinc and manganese from pyrometallurgy sludge or residuesInfo
- Publication number
- EP3172348A1 EP3172348A1 EP15809510.9A EP15809510A EP3172348A1 EP 3172348 A1 EP3172348 A1 EP 3172348A1 EP 15809510 A EP15809510 A EP 15809510A EP 3172348 A1 EP3172348 A1 EP 3172348A1
- Authority
- EP
- European Patent Office
- Prior art keywords
- leaching
- precipitation
- solution
- matrix
- enriched
- 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.)
- Withdrawn
Links
- 239000011701 zinc Substances 0.000 title claims abstract description 187
- 239000011572 manganese Substances 0.000 title claims abstract description 145
- 229910052748 manganese Inorganic materials 0.000 title claims abstract description 61
- 229910052725 zinc Inorganic materials 0.000 title claims abstract description 58
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 title claims abstract description 37
- PWHULOQIROXLJO-UHFFFAOYSA-N Manganese Chemical compound [Mn] PWHULOQIROXLJO-UHFFFAOYSA-N 0.000 title claims abstract description 30
- 238000011084 recovery Methods 0.000 title abstract description 6
- 239000010802 sludge Substances 0.000 title description 25
- 238000009853 pyrometallurgy Methods 0.000 title description 4
- 238000002386 leaching Methods 0.000 claims abstract description 168
- 238000001556 precipitation Methods 0.000 claims abstract description 121
- 238000000034 method Methods 0.000 claims abstract description 101
- 239000011133 lead Substances 0.000 claims abstract description 64
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims abstract description 50
- 229910052782 aluminium Inorganic materials 0.000 claims abstract description 25
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 claims abstract description 22
- 229910052745 lead Inorganic materials 0.000 claims abstract description 21
- 229910052742 iron Inorganic materials 0.000 claims abstract description 15
- 239000002699 waste material Substances 0.000 claims abstract description 10
- 239000000243 solution Substances 0.000 claims description 88
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 54
- 229910052751 metal Inorganic materials 0.000 claims description 49
- 239000002184 metal Substances 0.000 claims description 49
- 239000011159 matrix material Substances 0.000 claims description 46
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 42
- 239000002244 precipitate Substances 0.000 claims description 26
- 239000000203 mixture Substances 0.000 claims description 21
- 150000002739 metals Chemical class 0.000 claims description 20
- 239000000428 dust Substances 0.000 claims description 18
- 239000012535 impurity Substances 0.000 claims description 18
- 238000000926 separation method Methods 0.000 claims description 17
- BPQQTUXANYXVAA-UHFFFAOYSA-N Orthosilicate Chemical compound [O-][Si]([O-])([O-])[O-] BPQQTUXANYXVAA-UHFFFAOYSA-N 0.000 claims description 16
- 239000007788 liquid Substances 0.000 claims description 15
- 150000001875 compounds Chemical class 0.000 claims description 14
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 claims description 13
- 229910052979 sodium sulfide Inorganic materials 0.000 claims description 12
- 229910000016 manganese(II) carbonate Inorganic materials 0.000 claims description 11
- BVKZGUZCCUSVTD-UHFFFAOYSA-L Carbonate Chemical compound [O-]C([O-])=O BVKZGUZCCUSVTD-UHFFFAOYSA-L 0.000 claims description 10
- 229910000831 Steel Inorganic materials 0.000 claims description 10
- 239000011734 sodium Substances 0.000 claims description 10
- 239000010959 steel Substances 0.000 claims description 10
- JBQYATWDVHIOAR-UHFFFAOYSA-N tellanylidenegermanium Chemical compound [Te]=[Ge] JBQYATWDVHIOAR-UHFFFAOYSA-N 0.000 claims description 10
- 239000003638 chemical reducing agent Substances 0.000 claims description 9
- 150000004679 hydroxides Chemical class 0.000 claims description 8
- 229910000029 sodium carbonate Inorganic materials 0.000 claims description 8
- 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 7
- 239000012141 concentrate Substances 0.000 claims description 7
- 230000001376 precipitating effect Effects 0.000 claims description 7
- 239000000047 product Substances 0.000 claims description 7
- 229910052708 sodium Inorganic materials 0.000 claims description 7
- OYPRJOBELJOOCE-UHFFFAOYSA-N Calcium Chemical compound [Ca] OYPRJOBELJOOCE-UHFFFAOYSA-N 0.000 claims description 6
- 239000003929 acidic solution Substances 0.000 claims description 6
- 239000011575 calcium Substances 0.000 claims description 6
- 229910052791 calcium Inorganic materials 0.000 claims description 6
- 150000004649 carbonic acid derivatives Chemical class 0.000 claims description 6
- XLYOFNOQVPJJNP-UHFFFAOYSA-M hydroxide Chemical compound [OH-] XLYOFNOQVPJJNP-UHFFFAOYSA-M 0.000 claims description 6
- -1 carbonate compound Chemical class 0.000 claims description 5
- 238000009713 electroplating Methods 0.000 claims description 5
- ZLMJMSJWJFRBEC-UHFFFAOYSA-N Potassium Chemical compound [K] ZLMJMSJWJFRBEC-UHFFFAOYSA-N 0.000 claims description 4
- 239000003513 alkali Substances 0.000 claims description 4
- 239000006227 byproduct Substances 0.000 claims description 4
- AXCZMVOFGPJBDE-UHFFFAOYSA-L calcium dihydroxide Chemical compound [OH-].[OH-].[Ca+2] AXCZMVOFGPJBDE-UHFFFAOYSA-L 0.000 claims description 4
- 239000000920 calcium hydroxide Substances 0.000 claims description 4
- 229910001861 calcium hydroxide Inorganic materials 0.000 claims description 4
- 150000001768 cations Chemical class 0.000 claims description 4
- 150000001805 chlorine compounds Chemical class 0.000 claims description 4
- 239000011591 potassium Substances 0.000 claims description 4
- 229910052700 potassium Inorganic materials 0.000 claims description 4
- 239000000843 powder Substances 0.000 claims description 3
- 238000002156 mixing Methods 0.000 claims description 2
- 238000004064 recycling Methods 0.000 claims description 2
- 239000012487 rinsing solution Substances 0.000 claims description 2
- VDQVEACBQKUUSU-UHFFFAOYSA-M disodium;sulfanide Chemical compound [Na+].[Na+].[SH-] VDQVEACBQKUUSU-UHFFFAOYSA-M 0.000 claims 2
- 239000002002 slurry Substances 0.000 abstract description 5
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 abstract 2
- 239000000377 silicon dioxide Substances 0.000 abstract 1
- XLOMVQKBTHCTTD-UHFFFAOYSA-N Zinc monoxide Chemical compound [Zn]=O XLOMVQKBTHCTTD-UHFFFAOYSA-N 0.000 description 18
- XMWCXZJXESXBBY-UHFFFAOYSA-L manganese(ii) carbonate Chemical compound [Mn+2].[O-]C([O-])=O XMWCXZJXESXBBY-UHFFFAOYSA-L 0.000 description 11
- GRVFOGOEDUUMBP-UHFFFAOYSA-N sodium sulfide (anhydrous) Chemical compound [Na+].[Na+].[S-2] GRVFOGOEDUUMBP-UHFFFAOYSA-N 0.000 description 9
- 239000011787 zinc oxide Substances 0.000 description 9
- 238000000605 extraction Methods 0.000 description 8
- 238000000746 purification Methods 0.000 description 8
- 150000004763 sulfides Chemical class 0.000 description 8
- 239000000463 material Substances 0.000 description 7
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 6
- MUBZPKHOEPUJKR-UHFFFAOYSA-N Oxalic acid Chemical compound OC(=O)C(O)=O MUBZPKHOEPUJKR-UHFFFAOYSA-N 0.000 description 6
- 239000003795 chemical substances by application Substances 0.000 description 6
- 238000009854 hydrometallurgy Methods 0.000 description 6
- 239000011656 manganese carbonate Substances 0.000 description 6
- 235000006748 manganese carbonate Nutrition 0.000 description 6
- 229940093474 manganese carbonate Drugs 0.000 description 6
- NUJOXMJBOLGQSY-UHFFFAOYSA-N manganese dioxide Inorganic materials O=[Mn]=O NUJOXMJBOLGQSY-UHFFFAOYSA-N 0.000 description 6
- 239000007787 solid Substances 0.000 description 6
- WPBNNNQJVZRUHP-UHFFFAOYSA-L manganese(2+);methyl n-[[2-(methoxycarbonylcarbamothioylamino)phenyl]carbamothioyl]carbamate;n-[2-(sulfidocarbothioylamino)ethyl]carbamodithioate Chemical compound [Mn+2].[S-]C(=S)NCCNC([S-])=S.COC(=O)NC(=S)NC1=CC=CC=C1NC(=S)NC(=O)OC WPBNNNQJVZRUHP-UHFFFAOYSA-L 0.000 description 5
- 239000002562 thickening agent Substances 0.000 description 5
- MHAJPDPJQMAIIY-UHFFFAOYSA-N Hydrogen peroxide Chemical compound OO MHAJPDPJQMAIIY-UHFFFAOYSA-N 0.000 description 4
- RAHZWNYVWXNFOC-UHFFFAOYSA-N Sulphur dioxide Chemical compound O=S=O RAHZWNYVWXNFOC-UHFFFAOYSA-N 0.000 description 4
- TZCXTZWJZNENPQ-UHFFFAOYSA-L barium sulfate Chemical compound [Ba+2].[O-]S([O-])(=O)=O TZCXTZWJZNENPQ-UHFFFAOYSA-L 0.000 description 4
- 238000002354 inductively-coupled plasma atomic emission spectroscopy Methods 0.000 description 4
- 238000004519 manufacturing process Methods 0.000 description 4
- PXHVJJICTQNCMI-UHFFFAOYSA-N nickel Substances [Ni] PXHVJJICTQNCMI-UHFFFAOYSA-N 0.000 description 4
- 230000002378 acidificating effect Effects 0.000 description 3
- KRKNYBCHXYNGOX-UHFFFAOYSA-N citric acid Chemical compound OC(=O)CC(O)(C(O)=O)CC(O)=O KRKNYBCHXYNGOX-UHFFFAOYSA-N 0.000 description 3
- 239000010949 copper Substances 0.000 description 3
- 238000005868 electrolysis reaction Methods 0.000 description 3
- 239000000706 filtrate Substances 0.000 description 3
- 150000003839 salts Chemical class 0.000 description 3
- HRZFUMHJMZEROT-UHFFFAOYSA-L sodium disulfite Chemical compound [Na+].[Na+].[O-]S(=O)S([O-])(=O)=O HRZFUMHJMZEROT-UHFFFAOYSA-L 0.000 description 3
- 229940001584 sodium metabisulfite Drugs 0.000 description 3
- 235000010262 sodium metabisulphite Nutrition 0.000 description 3
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 3
- NLXLAEXVIDQMFP-UHFFFAOYSA-N Ammonia chloride Chemical compound [NH4+].[Cl-] NLXLAEXVIDQMFP-UHFFFAOYSA-N 0.000 description 2
- CIWBSHSKHKDKBQ-JLAZNSOCSA-N Ascorbic acid Chemical compound OC[C@H](O)[C@H]1OC(=O)C(O)=C1O CIWBSHSKHKDKBQ-JLAZNSOCSA-N 0.000 description 2
- 235000008733 Citrus aurantifolia Nutrition 0.000 description 2
- RWSOTUBLDIXVET-UHFFFAOYSA-N Dihydrogen sulfide Chemical compound S RWSOTUBLDIXVET-UHFFFAOYSA-N 0.000 description 2
- 241000196324 Embryophyta Species 0.000 description 2
- 235000011941 Tilia x europaea Nutrition 0.000 description 2
- 239000005083 Zinc sulfide Substances 0.000 description 2
- 229910052921 ammonium sulfate Inorganic materials 0.000 description 2
- 239000007864 aqueous solution Substances 0.000 description 2
- 238000010908 decantation Methods 0.000 description 2
- 239000011790 ferrous sulphate Substances 0.000 description 2
- 235000003891 ferrous sulphate Nutrition 0.000 description 2
- 238000001914 filtration Methods 0.000 description 2
- 229910000037 hydrogen sulfide Inorganic materials 0.000 description 2
- BAUYGSIQEAFULO-UHFFFAOYSA-L iron(2+) sulfate (anhydrous) Chemical compound [Fe+2].[O-]S([O-])(=O)=O BAUYGSIQEAFULO-UHFFFAOYSA-L 0.000 description 2
- 229910000015 iron(II) carbonate Inorganic materials 0.000 description 2
- 229910000359 iron(II) sulfate Inorganic materials 0.000 description 2
- 239000004571 lime Substances 0.000 description 2
- 229910052976 metal sulfide Inorganic materials 0.000 description 2
- 235000006408 oxalic acid Nutrition 0.000 description 2
- 238000007254 oxidation reaction Methods 0.000 description 2
- NIFIFKQPDTWWGU-UHFFFAOYSA-N pyrite Chemical compound [Fe+2].[S-][S-] NIFIFKQPDTWWGU-UHFFFAOYSA-N 0.000 description 2
- 229910052683 pyrite Inorganic materials 0.000 description 2
- 239000011028 pyrite Substances 0.000 description 2
- 230000001603 reducing effect Effects 0.000 description 2
- 230000000630 rising effect Effects 0.000 description 2
- 230000006641 stabilisation Effects 0.000 description 2
- 238000011105 stabilization Methods 0.000 description 2
- 229910001220 stainless steel Inorganic materials 0.000 description 2
- 239000010935 stainless steel Substances 0.000 description 2
- 239000000126 substance Substances 0.000 description 2
- DRDVZXDWVBGGMH-UHFFFAOYSA-N zinc;sulfide Chemical compound [S-2].[Zn+2] DRDVZXDWVBGGMH-UHFFFAOYSA-N 0.000 description 2
- 229910002703 Al K Inorganic materials 0.000 description 1
- 241000894006 Bacteria Species 0.000 description 1
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 description 1
- MYMOFIZGZYHOMD-UHFFFAOYSA-N Dioxygen Chemical compound O=O MYMOFIZGZYHOMD-UHFFFAOYSA-N 0.000 description 1
- 229910000616 Ferromanganese Inorganic materials 0.000 description 1
- WQZGKKKJIJFFOK-GASJEMHNSA-N Glucose Natural products OC[C@H]1OC(O)[C@H](O)[C@@H](O)[C@@H]1O WQZGKKKJIJFFOK-GASJEMHNSA-N 0.000 description 1
- GUBGYTABKSRVRQ-QKKXKWKRSA-N Lactose Natural products OC[C@H]1O[C@@H](O[C@H]2[C@H](O)[C@@H](O)C(O)O[C@@H]2CO)[C@H](O)[C@@H](O)[C@H]1O GUBGYTABKSRVRQ-QKKXKWKRSA-N 0.000 description 1
- FYYHWMGAXLPEAU-UHFFFAOYSA-N Magnesium Chemical compound [Mg] FYYHWMGAXLPEAU-UHFFFAOYSA-N 0.000 description 1
- PMZURENOXWZQFD-UHFFFAOYSA-L Sodium Sulfate Chemical compound [Na+].[Na+].[O-]S([O-])(=O)=O PMZURENOXWZQFD-UHFFFAOYSA-L 0.000 description 1
- 240000008042 Zea mays Species 0.000 description 1
- 235000005824 Zea mays ssp. parviglumis Nutrition 0.000 description 1
- 235000002017 Zea mays subsp mays Nutrition 0.000 description 1
- 235000019270 ammonium chloride Nutrition 0.000 description 1
- BFNBIHQBYMNNAN-UHFFFAOYSA-N ammonium sulfate Chemical compound N.N.OS(O)(=O)=O BFNBIHQBYMNNAN-UHFFFAOYSA-N 0.000 description 1
- 235000011130 ammonium sulphate Nutrition 0.000 description 1
- 239000011668 ascorbic acid Substances 0.000 description 1
- 235000010323 ascorbic acid Nutrition 0.000 description 1
- 229960005070 ascorbic acid Drugs 0.000 description 1
- BRPQOXSCLDDYGP-UHFFFAOYSA-N calcium oxide Chemical compound [O-2].[Ca+2] BRPQOXSCLDDYGP-UHFFFAOYSA-N 0.000 description 1
- 150000001720 carbohydrates Chemical class 0.000 description 1
- 235000014633 carbohydrates Nutrition 0.000 description 1
- 239000003153 chemical reaction reagent Substances 0.000 description 1
- 235000015165 citric acid Nutrition 0.000 description 1
- 229910017052 cobalt Inorganic materials 0.000 description 1
- 239000010941 cobalt Substances 0.000 description 1
- GUTLYIVDDKVIGB-UHFFFAOYSA-N cobalt atom Chemical compound [Co] GUTLYIVDDKVIGB-UHFFFAOYSA-N 0.000 description 1
- 239000000470 constituent Substances 0.000 description 1
- 229910052802 copper Inorganic materials 0.000 description 1
- 235000005822 corn Nutrition 0.000 description 1
- 229910001882 dioxygen Inorganic materials 0.000 description 1
- 238000010891 electric arc Methods 0.000 description 1
- 239000007789 gas Substances 0.000 description 1
- 239000008103 glucose Substances 0.000 description 1
- 159000000011 group IA salts Chemical class 0.000 description 1
- 231100001261 hazardous Toxicity 0.000 description 1
- 150000002500 ions Chemical class 0.000 description 1
- DALUDRGQOYMVLD-UHFFFAOYSA-N iron manganese Chemical compound [Mn].[Fe] DALUDRGQOYMVLD-UHFFFAOYSA-N 0.000 description 1
- 239000008101 lactose Substances 0.000 description 1
- 239000011777 magnesium Substances 0.000 description 1
- 229910052749 magnesium Inorganic materials 0.000 description 1
- VTHJTEIRLNZDEV-UHFFFAOYSA-L magnesium dihydroxide Chemical compound [OH-].[OH-].[Mg+2] VTHJTEIRLNZDEV-UHFFFAOYSA-L 0.000 description 1
- 239000000347 magnesium hydroxide Substances 0.000 description 1
- 229910001862 magnesium hydroxide Inorganic materials 0.000 description 1
- 238000005065 mining Methods 0.000 description 1
- 235000013379 molasses Nutrition 0.000 description 1
- 229910052759 nickel Inorganic materials 0.000 description 1
- 230000003647 oxidation Effects 0.000 description 1
- 238000006479 redox reaction Methods 0.000 description 1
- 239000010801 sewage sludge Substances 0.000 description 1
- 150000004760 silicates Chemical class 0.000 description 1
- 239000002893 slag Substances 0.000 description 1
- 229910052938 sodium sulfate Inorganic materials 0.000 description 1
- 235000011152 sodium sulphate Nutrition 0.000 description 1
- 239000004753 textile Substances 0.000 description 1
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
- C22B47/00—Obtaining manganese
-
- 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
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/30—Obtaining zinc or zinc oxide from metallic residues or scraps
-
- 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
-
- 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
-
- 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
- C22B7/007—Wet processes by acid leaching
-
- 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
Hydrometallurgical techniques for the selective recovery of Zn, Mn and Pb in that order from sludges or various metallurgical residues can include leaching steps to extract the zinc and manganese; the steps of removing iron, aluminum and silica from the leaching solutions; and recovering manganese and zinc by precipitation. Mn can be precipitated from an initial leaching solution as well as a secondary leaching solution obtained by leaching the slurry of the initial leaching step. Various other waste streams can also be treated.
Description
RECOVERY OF ZINC AND MANGANESE FROM PYROMETALLURGY SLUDGE OR
RESIDUES
FIELD
The technical field relates to the treatment of pyrometallurgical sludge or residues, and more particularly to the recovery of components from such materials.
BACKGROUND
Among the numerous by-products generated by the steel industry, sewage sludge of smokes of blast furnaces can be very rich in manganese, zinc and lead (between 5 and 40%). These materials have generally been deposited in periphery of the sites of steel industry.
Zn is an important metal in metallurgical, chemical and textile industries. It mainly extracted from sulphide ores including Zn. Portions of Zn can be recovered from secondary sources such as electric furnace dust containing different levels of impurities depending on the source (Jha et al. 2001). Pyrometallurgical and hydrometallurgical processes are usually used to recover Zn in the ore or from secondary sources. Nevertheless, the pyrometallurgical processes consume significant energy and require process gas treatment during Zn recovery. The majority of studies have leached zinc with sodium hydroxide (Charpentier et al., 2007), ammonium chloride (Olper et al., 1993) and sulfuric acid (Dvorak et al., 2005).
Pyrolusite (Mn02) is stable in acidic and alkaline conditions. Extraction of Mn (IV) must be carried out under reducing conditions (Naik et al. 2000). Various techniques for leaching ores or nodules have been studied for the last 20 years. Most studies function with sulfuric acid as a reducing agent, such as sulfuric acid leaching, coupled hydrogen peroxide (Allen et al, 1988; Jiang et al, 2004), to oxalic acid (Sahoo et al, 2001.), ferrous sulfate (Brantley et al, 1968), pyrite (Vracar et al, 2000) and aqueous sulfur dioxide (S02) Chow et al, 2012b, Ward (2005a), and Ward (2005b).
Leaching with hydrochloric acid (HCI) is also used with Ni (Chen et al, 1992), pyrite Kanungo (1999a) and hydrogen peroxide (El Hazek et al, 2006). Sayilgan et al. (2010)
conducted a series of studies of Mn ore leaching with hydrochloric acid or sulfuric acid in the presence of a reducing agent such as ascorbic acid, citric acid and oxalic acid.
Other studies have shown reducing properties using acidic carbohydrates waste from agriculture, renewable and non-hazardous such as glucose (Furnali et al 2006) molasses (Lasheen et al, 2009) lactose (Veglio et al, 2000), corn cobs (Tian et al, 2010) or sawdust (Hariprasad et al, 2007). Some authors use bacteria for bioleaching, as in Veglio et al. (1997).
Some studies have looked at hydrometallurgical process including the steps of leaching, extraction and purification of Mn in the form of oxides or salts. These methods mainly use reagents such as ferrous sulfate and S02 in sulfuric acid. Those chemicals are generally appreciated for their low cost and their effectiveness (see Chow et al., 2012 WO2012024294).
In a hydrometallurgical process, the principle of selective leaching is rarely used. Solution can contain impurities that can affect the production of pure Mn compounds. Purification is usually required to remove impurities present in aqueous solutions before proceeding to the recovery of the metal. The precipitation of the metals as hydroxides or sulfides are the main purification techniques used to recover the Mn and Zn in the solution. One challenge is to precipitate unwanted metals and to keep in solution Zn and Mn with a cost as low as possible.
Precipitation of metals as hydroxides is a common method for removing metals from a solution in hydrometallurgical processes (Zhang et al., 2007).
After precipitation of Fe3+ and Al3+ ions in the form of hydroxides, precipitation of the remaining impurities as sulphides is generally employed. This extraction is based on the difference in solubility between the metal sulfides as function of pH and temperature. Dihydrogen sulfide (H2S), sodium sulfate (Na2S), barium sulfate (BaS), or ammonium sulfate (NH4)2S04 are generally used for the precipitation of impurities in the form of metal sulfides. Pakarinen et al. (2011) use a precipitation with Na2C03 under 02 supply to allow the separation of Fe2+ and Mn2+ ions in solution. At pH 6 and under ORP > 200mV, it happens to precipitate the Fe2 + in FeC03 keeping the Mn2+ in solution. The FeC03 was then removed by filtration. The use of carbonate allows a good separation of Fe and Mn with an optimal price/quality ratio. Cementation is also a technique commonly
used in hydrometallurgy. This technique includes the precipitation of a metal from an aqueous solution of a salt thereof with another metal. This redox reaction connects a more electropositive metal which is deposited on a less electropositive metal from solution. This technique is used in zinc hydrometallurgy to precipitate the metals present in the solution as Cu by adding Zn powder (Rizet et al, 2000).
There are various challenges in the treatment of pyrometallurgical sludge or residue, particularly for extracting components such as Zn and Mn.
SUMMARY
Techniques described herein relate selective extraction of zinc and manganese from metal sludge or dust by utilizing acidic leaching (e.g., using sulfuric acid) and generating a high-concentration lead slag. Various implementations are well suited to treating pyrometallurgic sludge or residue, and particularly to recover valuable constituents from sludge or residue, such manganese, zinc and lead in particular
In some implementations, the techniques enable steel residues to be treated by selectively and sequentially leaching the zinc and manganese, followed by effective purification to produce marketable product of each these metals.
Currently there is no method adapted to decontaminate and recover Zn, Mn and Pb contents in the dust generated by the ferromanganese industry. These residues have historically been deposited in basins bordering steel production plants, because legislation and metal prices over the last 100 years the industry did not promote the treatment of these waste streams. Given the market value of the metals and their general shortage, these brownfield sites can be considered as secondary resources. Techniques capable of profitably exploiting this type of residue (sludge), in order to produce compounds of Mn, Pb and Zn pure enough to be recoverable, can be advantageous for treating such wastes and also producing valuable products. In some implementations, the techniques include selectively leaching Zn, Mn and Pb in order to obtain a concentrate of Mn, a concentrate of Zn and residue rich in Pb.
Various tests have identified leaching agents and effective and affordable discount. Further steps include the purification of each leachate by removing unwanted metals such as iron (Fe), cobalt (Co), nickel (Ni), copper (Cu) and aluminum (Al), and then
production of a marketable product of each valuable metal (e.g., Zn, Mn and Pb). In some scenarios, the techniques lead to the production of salts of Mn, Zn free from substantial impurities, as well as a final residue rich enough in Pb to be recycled.
In some implementations, there is provided a process for treating a matrix comprising zinc (Zn), manganese (Mn) and lead (Pb), comprising: leaching the matrix to generate a leaching solution enriched in Zn and a fraction enriched in Mn and Pb; subjecting the leaching solution to Zn-precipitation to produce Zn-based precipitates and a Zn-depleted fraction comprising Mn compounds; leaching the fraction enriched in Mn and Pb to produce a Mn-enriched leaching solution and a Pb-enriched fraction; and subjecting the Zn-depleted fraction and/or the Mn-enriched leaching solution to Mn-precipitation to produce Mn-based precipitates.
In some implementations, the leaching of the matrix is performed using an aqueous acidic solution.
In some implementations, the aqueous acidic solution comprises H2S04.
In some implementations, the leaching of the matrix comprises multiple leaching runs with the aqueous acidic solution and multiple rinsing runs with an aqueous rinsing solution.
In some implementations, the leaching of the matrix comprises at least three leaching runs and/or at least three rinsing runs.
In some implementations, the process also includes, prior to the Zn-precipitation, removing additional metal components from the leaching solution enriched in Zn. In some implementations, the additional metal components include aluminum (Al). In some implementations, the additional metal components include iron (Fe).
In some implementations, the removing of the additional metal components comprises precipitating the additional metal components. In some implementations, precipitating of
the additional metal components comprises adding a hydroxide of a divalent cation. In some implementations, the hydroxide of a divalent cation comprises Ca(OH)2 or NaOH.
In some implementations, the Zn-precipitation comprises adding NaOH or Na2S to produce ZnO or ZnS as the Zn-based precipitates.
In some implementations, the process also includes, prior to the Mn-precipitation, removing additional metallic impurities from the Zn-depleted fraction and/or the Mn- enriched leaching solution, including aluminum, iron, lead, zinc or a combination thereof. In some implementations, precipitating of the additional metal components prior to the Mn-precipitation comprises adding NaOH and Na2S to the Zn-depleted fraction and/or the Mn-enriched leaching solution.
In some implementations, the Mn-precipitation comprises adding a carbonate compound to produce Mn-carbonates as the Mn-based precipitates. In some implementations, the carbonate compound comprises Na2C03 to produce MnC03.
In some implementations, the process also includes, prior to the leaching of the matrix, a step of determining a composition of the matrix. In some implementations, determining the composition of the matrix comprises determining the concentrations of Zn and Mn. In some implementations, determining the composition of the matrix comprises determining the concentration of Pb. In some implementations, determining the composition of the matrix includes determining the concentrations of Al and/or Fe.
In some implementations, the process also includes treating the Pb-enriched fraction to recover Pb therefrom.
In some implementations, the process also includes subjecting the Zn-depleted fraction to Mn-precipitation in a first reactor to produce a first stream of Mn-based precipitates; and subjecting the the Mn-enriched leaching solution to Mn-precipitation in a second reactor to produce a second stream of Mn-based precipitates.
In some implementations, the process further includes producing a Mn-depleted stream in the second reactor; and subjecting at least a portion of the Mn-depleted stream to further Mn-precipitation. In some implementations, the further Mn-precipitation of the Mn- depleted stream is conducted in the first reactor.
In some implementations, the process further includes recycling a portion of the Mn- depleted stream from the second reactor for addition to the matrix before or during the leaching thereof.
In some implementations, the process further includes recovering a metals-depleted liquid from the Mn-precipitation in the first reactor and adding at least a portion of the metals-depleted liquid to the matrix before or during the leaching thereof.
In some implementations, the leaching of the matrix is performed in atmospheric pressure conditions.
In some implementations, one or more of the leaching steps are performed in a counter- current reactor.
In some implementations, one or more of the leaching steps are performed in an agitated tank reactor.
In some implementations, one or more of the leaching steps are performed in a batch reactor or a continuous.
It is noted that the leaching reactors can use recirculated streams derived from other units of the process. For instance, the recirculated streams can include leaching solutions from which one or more metals (e.g., valuable metal components) have been removed, such as streams that have been subjected to Mn-precipitation. It should be noted that the recirculation can be done at various points in the overall process, using different recycle streams depending on various factors, such as economic considerations, make-up water availability, build-up of certain components in the recycled streams, and so on.
In some implementations, the matrix comprises a powder or dust. In some implementations, the matrix comprises a by-product or waste stream derived from metals processing. In some implementations, the matrix comprises steel plant dust.
In some implementations, there is provided a process for treating a matrix comprising zinc (Zn), manganese (Mn) and lead (Pb), comprising selective and sequential leaching of Zn, Mn and Pb in order to obtain a Mn concentrate, a Zn concentrate, and a Pb- enriched residue.
In some implementations, there is provided a method for the hydrometallurgical treatment of a matrix comprising manganese (Mn), zinc (Zn), lead (Pb), iron (Fe), aluminum (Al), and at least one of calcium, chlorides, sodium and potassium as impurities, the method comprising: a first leaching step comprising: leaching the matrix at atmospheric pressure in a reactor using a first leaching solution to product a first leach liquor and a first leach residue, the first leaching solution comprising a sulfuric acid solution; and separating the first leach liquor from the first leach residue after the first leaching; a second leaching step comprising: leaching the first leach residue at atmospheric pressure using a second leaching solution to form a second leach liquor and a second leach residue, the second leaching solution comprising sulfuric acid and a reducing agent; and separating the second leach liquor from the second leach residue; a Mn-precipitation step wherein Mn compounds are precipitated from the first and second leach liquors, including contacting the liquors or streams derived therefrom, with a carbonate at a pH to precipitate Mn carbonate compounds.
In some implementations, the process also includes selecting a quantity of the carbonate to adjust the pH of the first and second leach liquor sufficiently to precipitate out Mn carbonate compounds.
In some implementations, the pH of the second leaching is between about 3 and about 5.
In some implementations, the reducing agent used in the second leaching step comprises Na2S205.
In some implementations, the process further includes subjecting the first liquor to Zn- precipitation prior to Mn-precipitation.
In some implementations, the Zn-precipitation includes electroplating Zn as Zn metal.
In some implementations, the Zn-precipitation includes forming Zn sulfide and/or Zn oxide precipitates that are precipitated out of the solution.
In some implementations, the process includes separating the Zn sulfide and/or Zn oxide precipitates from the solution to form a Zn sulfide and/or Zn oxide containing residue, and a Zn-depleted first liquor that is subjected to the Mn-precipitation step.
In some implementations, the Zn-precipitation comprises adding alkali or sulfide to the first liquor to precipitate Zn compounds; and then preforming a solid-liquid separation to produce a Zn sulfide and/or oxide cake.
In some implementations, the process includes subjecting the first liquor to a Fe, silicate and Al precipitation step to form Fe, silicate and Al hydroxide compounds and a depleted liquor; separating the Fe, silicate and Al hydroxide compounds from the depleted liquor; and subjecting the depleted liquor to the Zn-precipitation.
In some implementations, the Fe, silicate and Al precipitation step comprises adding alkali to increase the pH between about 4 and about 5; and then preforming a solid-liquid separation to produce a mixed Fe, silicate and Al hydroxide cake.
In some implementations, the Mn-precipitation step comprises adding the carbonate to the first and second liquors to increase the pH to between about 7.5 and about 8.5; and then performing a solid-liquid separation to produce a Mn-carbonate cake.
In some implementations, the second leach residue contains between 10 wt% and 30 wt% of Pb.
In some implementations, a leaching period of mixing the sulfuric acid solution for the first and/or second leaching steps is between about 10 minutes and about 24 h.
In some implementations, the temperature of the sulfuric acid solution is between about 20°C and about 40°C.
In some implementations, the process includes one or more features of claims 1 to 34.
In some implementations, there is provided a system for treating a matrix comprising zinc (Zn) and manganese (Mn), comprising: a leaching unit for leaching the matrix to generate a Zn-enriched leaching solution, and a fraction enriched in Mn and Pb; a Zn-precipitation unit for subjecting the Zn-enriched leaching solution to Zn- precipitation to produce Zn-based precipitates and a Zn-depleted fraction comprising Mn compounds; and a leaching unit for leaching the fraction enriched in Mn and Pb to produce a Mn- enriched leaching solution and a Pb-enriched fraction; and a Mn-precipitation system for subjecting the Zn-depleted fraction and/or the Mn- enriched leaching solution, or streams derived therefrom, to Mn-precipitation to produce Mn-based precipitates.
In some implementations, the Mn-precipitation system comprises a first Mn-precipitation unit for subjecting the Zn-depleted fraction to Mn-precipitation; and a second Mn- precipitation unit for subjecting the Mn-enriched leaching solution to Mn-precipitation.
In some implementations, the systems includes an additional precipitation unit for receiving the Zn-enriched leaching solution and removing at least one of aluminum (Al), iron (Fe) and silicate therefrom, to produce a liquor for supplying to the Zn-precipitation unit.
In some implementations, there is provided a system for the hydrometallurgical treatment of a matrix comprising manganese (Mn), zinc (Zn), lead (Pb), iron (Fe), aluminum (Al), and at least one of calcium, chlorides, sodium and potassium as impurities, the system comprising: a first leaching unit for leaching the matrix at atmospheric pressure in a reactor using a first leaching solution to product a first leach liquor and a first leach residue, the first leaching solution comprising a sulfuric acid solution;
a first separation unit for separating the first leach liquor from the first leach residue after the first leaching; a second leaching unit for leaching the first leach residue at atmospheric pressure using a second leaching solution to form a second leach liquor and a second leach residue, the second leaching solution comprising sulfuric acid and a reducing agent; a second separation unit for separating the second leach liquor from the second leach residue; and a Mn-precipitation system for precipitating Mn compounds from the first and second leach liquors, or streams derived therefrom, by contacting with a carbonate at a pH to precipitate Mn carbonate compounds.
In some implementations, the Mn-precipitation system comprises a first Mn-precipitation unit for subjecting the first leach liquor to Mn-precipitation; and a second Mn-precipitation unit for subjecting the second leach liquor to Mn-precipitation.
In some implementations, the system includes an additional precipitation unit for receiving the first leach liquor and removing at least one of aluminum (Al), iron (Fe) and silicate therefrom, to produce a metals depleted liquor comprising Zn and Mn; and a Zn- precipitation unit for receiving the metals depleted liquor and producing Zn precipitates and a Zn-depleted fraction, wherein the first precipitation unit is configured to receive the Zn-depleted fraction to be subjected to the Mn-precipitation.
In some implementations, the system also includes one or more operational or structural features of the processes described in the description or drawings.
Various implementations, aspects and features of the techniques will be described below and illustrated in the drawings.
BRIEF DESCRIPTION OF THE DRAWINGS
Figure 1 is a graph showing Zn concentration (mg/L) during electrolysis.
Figure 2 is a diagrammatic representation of a flow sheet depicting a method for the leaching of a metallic sludge.
Figure 3 is a diagrammatic representation of a process for treating a metallic sludge.
Figure 4 is a diagrammatic representation of another process for treating a metallic sludge.
DETAILED DESCRIPTION
Various implementations of techniques will be described for treating metallic sludges for recovering components, such as Zn and Mn.
Referring to Figures 3 and 4, in some implementations the process begins with metallic sludge 10, which may be obtained from "brownfield" or elsewhere. The metallic sludge 10 can contain between 1 and 40 percent of Zn, between 0.5 and 28 % of Mn and between 1 and 20 % of Pb, for example. It should be noted that the source material that is treated can be a mixture of metallic sludge streams from multiple sources and/or a pre-treated metallic sludge stream or another stream derived from metallic sludge.
In addition to manganese, zinc and lead, such metallic sludges may contain at least one of the following elements: iron, aluminum, calcium, silicates, sodium and magnesium.
In some implementations, the metallic sludge 10 is reacted at ambient temperature with sulfuric acid at step 1 1 , referred to as Zn leaching step. The resource material pulp density can be between 5 and 15% by weight.
Optionally, there can be multiple (e.g., three) steps of leaching with a solution of sulfuric acid with a concentration of 0.25 M and multiple (e.g., three) steps of rinsing. Aqueous sulfuric acid reacts with zinc oxide according the following equation:
Equation 1 : ZnO + H2S04 -> Zn2+ + H20 + S04 2"
H2S04 reacts with divalent manganese (MnO) according the following equation:
Equation 2: MnO + H2S04 Mn2+ + H20 + S04 2"
In some implementations, the removal of impurities from leach liquor ("LL" which may also be referred to as pregnant leach solution "PLS") can be accomplished in one stage. For example, in some implementations, the pH of the solution is changed from about 4.25 to about 4.75 to precipitate silicate, iron and aluminum at a mix tank, shown at step
14, referred to as Al and Fe precipitation step. This pH may be achieved in various way, including the addition to the solution of alkaline earth hydroxides or alkaline salt, such as NaOH, Mg(OH)2 or Ca(OH)2. The precipitated solid can be separated from the treated leach liquor in a thickener or a filter in a solid-liquid separation step 15.
In some implementations, the pH of the solution can then be raised; for example, sodium sulfide or Na2S can be added to the solution to precipitate zinc as sulfides, the pH of the leach liquor can be about 4.25 to about 4.75. Zinc can also be precipitated and recovered as zinc oxide with a stoichiometric addition of sodium hydroxide, the pH of the leach liquor would be changed to about 7.0 to about 10.0, at step 16, referred to as Zn precipitation step. Zinc can also be recovered through electroplating with specific conditions. The sulphides or oxides precipitates can be separated from the treated leach liquor in a thickener or a filter in another solid-liquid separation step 17.
Optionally, the pH of the solution can then be raised to about 8.5 with Na2C03 to precipitate and recover manganese carbonate, at step 18, referred to as Mn precipitation step The solid manganous carbonate is separated from the solution by a thickener or a filter, in another solid-liquid separation step 19.
In some implementations, after zinc leaching, the leached residue (from the first step 11 of Zn leaching) can be reacted in a stirred tank at ambient temperature with sulfuric acid and sodium metabisulfite to reduce manganese dioxide, at step 12, referred to as Mn leaching step. Optionally, there can be 2 steps of leaching and 2 steps of rinsing.
Sodium metabisulfite reacts with Mn02 according the following equation:
Equation 3: 2Mn02 + H2S04 + Na2S205 -> 2Mn2+ + H20 + 3S04 2" + 2Na+
In some implementations, the pH of the solution can be between about 3.0 and about 5.0 to leach manganese without leaching iron or aluminum.
In some implementations, additional removal of impurities from leach can be accomplished in one stage. For example, referring to Figure 4, the pH of the solution can be changed from about 4.80 to about 5.60 to precipitate aluminum, iron, lead and zinc in a mix tank, shown in step 21 , referred to as Al, Fe, Pb and Zn precipitation step. Achievement of the desired pH may include the addition to the solution of sodium hydroxide and sulfide, such Na2S, to precipitate impurities as sulfides. The precipitated
solid can be separated from the treated leach liquor in a thickener or a filter in a solid- liquid separation step 22.
Referring to Figures 3 and 4, in some implementations, after filtration the pH of the solution is then raised to about 8.5 with Na2C03 to precipitate and recover manganese carbonate, at step 23, referred to as another Mn precipitation step. The solid manganous carbonate can be separated from the solution by a thickener or a filter in another solid- liquid separation step 24.
Final residues can contain a high proportion of lead, between 10 and 30% for example, which can be revalorize by hydrometallurgical and/or pyrometallurgical methods.
EXAMPLE I: LEACHING
A leaching test was performed using steel dust. The composition of the dust is given in Table 1. In the test, 400 g of dust was lixiviated (i.e., leached) with 4 liters of water (10% pulp density by weight) in a stirred plastic reactor with baffle.
Table 1 : Elements concentration in the metallic sludge (g/kg)
Mn Pb Al K Ca Fe Mg Sn Zn S
198 72.9 45.3 6.88 56.3 52.3 1 1 .5 13.1 18.7 3.23
After the leaching test for zinc, the slurry was filtered. A small portion was dried, crushed, digested and analyzed with ICP AES is given in Table 2. A sample of the filtrate was also analyzed.
Table 2: Elements concentration in the metallic sludge (g/kg)
Mn Pb ΑΪ K Ca Fe Mg Sn Zn S 231 102 ΪΟΪ 055 402 567 i~47 A ΣΪ3 55.5
It is noted that by operating in a pH range of 1.5 to 3.5, the sulfuric acid concentration used for each leaching is 0.25 M, about 97% of the zinc and 13% of manganese is extracted. These results shows that as Mn is mainly found here as Mn02 the first
leaching step preferentially leaches Zn and not much Mn or Pb. Leaching times did not exceed 30 minutes and rinsing times did not exceed 10 minutes. The weight of dry residue after leaching was 252 g. After the zinc leaching, a second leaching step included leaching the first leach residue in a second leaching solution with sulfuric acid and reductant such as sodium metabisulfite. The elements concentration in the leach liquors are shown in the Table 3.
Table 3: Elements concentration in leach liquor (mg/L)
Volume [H2S04]
PH Mn Pb Al K Ca Fe Mg Sn Zn S (mL) (mol/L)
L1 Zn 3054 0.25 3.48 439 4.78 530 375 519 2.54 512 0.81 463 2882
L2 Zn 4168 0.25 2.49 967 7.16 1773 236 51 1 69.4 402 41 .2 803 5915
L3 Zn 4176 0.25 1 .46 740 6.30 988 95.6 593 101 1 176 40.1 407 7466
R1 Zn 4000 0.0 1 .74 272 4.68 353 33.7 562 21 1 59.6 12.4 134 2564
R2 Zn 4000 0.0 2.1 1 104 3.70 1 15 13.9 564 29.9 18.6 2.70 42.8 1 171
R3 Zn 4000 0.0 2.40 51 .1 3.83 40.7 7.87 593 6.58 6.31 0.68 16.4 824
Extraction
12.8 0.41 83.7 100 58.0 26.3 93.7 7.72 96.7 -
%
L1Mn 2830 0.25 4.76 1 1998 9.90 0.00 13.5 573 0.00 49.3 12.0 44.3 17971
L2 Mn 2590 0.10 3.65 7954 9.04 155 8.25 548 5.44 45.2 6.45 93.4 14228
R1 Mn 2540 0.0 3.77 636 5.38 24.5 2.15 542 0.27 4.46 1 .87 9.34 1561
R2 Mn 2540 0.0 3.79 2226 5.94 56.7 1 .65 517 1 .15 14.0 1 .85 28.3 4263
Extraction
90.9 0.70 87.1 100 83.5 26.4 100 8.87 100 -
%
After the second leaching, the slurry was filtered. A small portion was dried, crushed, digested and analyzed with ICP AES is given in Table 4. A sample of the filtrate was also analyzed.
Table 4: Elements concentration in the metallic sludge (g/kg)
Mn Pb Al K Ca Fe Mg Sn Zn S
24.2 171 12.6 39.0 18.5 94.5 0.83 26.4 1 .34 61 .8
It is noted that by operating in a pH range 3.5 and 4.5, about 90.9 % of the manganese is extracted. Leaching times did not exceed 120 minutes and rising times did not exceed 10 minutes. The weight of dry residue after leaching is 138 g. This final residue contained 17.1 % Pb and is recyclable by other processes.
EXAMPLE II: LEACHING
Another leaching test was performed using steel dust more concentrated in zinc and less in manganese. The composition of the dust is given in Table 5. In the test, 20 g of dust was lixiviated with 200 mL of a solution with a sulfuric acid concentration of 0.25 mol/L (10% pulp density by weight) in an erlenmeyer with baffle. Table 6 presents the elements concentration in each leaching solution.
Table 5: Elements concentration in the metallic sludge (g/kg)
Mn Pb Al K Ca Fe Mg Sn Zn S
0.74 108 19.8 0.0 10.2 62.3 2.03 0.54 392 15.9
After the leaching test of zinc, the slurry was filtered. A small portion was dried, crushed, digested and analyzed with ICP AES is given in Table 7. A sample of the filtrate was also analyzed. The sulfuric acid concentration used for each leaching is 0.25 mol/L, about 89.7% of the zinc is extracted. Times of leaching did not exceeded 30 minutes and times of rising did not exceeded 10 minutes. The weight of dry residue after leaching is 7.4 g. It is not necessary to do manganese leaching because the concentration of manganese in the metallic sludge is too low. This final residue contains 27.4% Pb and is recyclable by other processes.
Table 6: Elements concentration in leach liquor (mg/L)
(mol/L)
L1 Zn 190 0.25 5.66 8.33 13.0 18.0 248 23.5 33.1 0.0 10817
L2 Zn 210 0.25 5.36 7.42 8.75 30.2 138 43.3 27.9 0.0 12613
L3 Zn 226 0.25 2.89 13.2 1 1.5 1134 83.7 671 29.2 40.0 9445
R1 Zn 200 0.0 3.80 1 .89 5.91 81.0 11 .4 63.4 2.90 0.0 946
Extraction
44.1 0.4 71.4 47.7 14.3 47.2 83.5 89.7
%
Table 7: Elements concentration in the metallic sludge (g/kg)
Mn Pb Al K Ca Fe Mg Sn Zn S
1 .0 274 24.0 1.0 16.1 136 4.0 - 39.7 24.0
EXAMPLE III: LEACHING
Another leaching test was performed using steel dust more concentrates in manganese and less in zinc and lead. The composition of the dust is given in Table 8. In the test, 20 g of dust was lixiviated with 200 mL of a solution with a sulfuric acid (10% pulp density by weight) in an erlenmeyer with baffle. Table 9 presents the element concentration in each leaching solution.
Table 8: Elements concentration in the metallic sludge (g/kg)
Mn Pb Al K Ca Fe Mg Sn Zn S
278 3.93 58.0 7.88 56.7 12.7 16.6 0.51 7.74 5.60
Table 9: Elements concentration in leach liquor (mg/L)
Volume
PH Mn Pb Al K Ca Fe Mg Sn Zn S
(mL)
L1 Zn 150 3.31 173 0.0 617 250 662 1 .50 652 0.0 99.4 2712
L2 Zn 204 1 .83 744 0.9 1712 213 603 17.2 492 2.80 299 5641
L3 Zn 204 1 .22 670 1.9 1 152 107 632 79.0 258 2.00 178 6928
R1 Zn 202 1 .56 200 0.0 541 45.0 591 8.00 11 1 0.60 67.2 2758
R2 Zn 202 1 .94 61.2 0.0 230 24.8 612 0.80 47.3 0.0 22.8 1367
R3 Zn 202 2.34 23.1 0.0 109 14.2 639 0.0 26.1 0.0 8.20 889
L1 Mn 176 3.81 12846 4.65 22.2 26.8 562 0.14 86.6 7.92 40.7 19464
L2 Mn 162 3.62 81 16 5.17 28.5 12.3 533 5.69 61.8 5.93 51 .5 14078
R1 Mn 150 3.66 3063 1.57 9.17 3.79 507 0.77 21.7 2.00 18.1 5162
R2 Mn 152 4.1 1 1230 5.89 8.56 2.86 531 0.00 7.54 0.0 6.92 2147
Extraction
82.6 4.3 74.7 80.9 93.9 8.9 95.7 36.8 97.8
%
After the leaching test of zinc and manganese, the slurry was filtered. A small portion was dried, crushed, digested and analyzed with ICP AES is given in Table 10. The weight of dry residue after leaching is 5.2 g. About 82.6% of manganese principally in the step of manganese leaching (L1 Mn to R2 Mn) and 97.8% of zinc is extracted principally in the steps of zinc leaching (L1 Zn to R3 Zn).
Table 10: Elements concentration in the metallic sludge (g/kg)
Mn Pb Al K Ca Fe Mg Sn Zn S
66.3 13.1 80.8 2.4 1 16 52.4 9.0 23.9 6.9 125
EXAMPLE IV: LEACH LIQUOR PURIFICATION
In this example, the leach liquor purification is performed using a technique such as precipitation. The aim of precipitation is to remove iron and aluminum from the leach liquor after the zinc leaching test. The compositions of the leach liquor and the initial pH are in Table 11. For each precipitation step, 1 liter of leach liquor has been used.
Table 11 : Elements concentration in the leach liquor (mg/L)
PH Mn Pb Al K Ca Fe Mg Sn Zn S Na
L1 Zn 3.17 439 4.8 530 375 519 2.50 512 0.8 463 2882 28.5
L2 Zn 2.42 967 7.2 1773 236 51 1 69.4 403 41.2 803 5915 11 .9
L3 Zn 1 .46 740 6.3 988 95.6 593 1011 176 40.1 406 7466 4.70
The pH of the solution is changed to about 4.25 and 4.75 to precipitate silicate, iron and aluminum in a mix tank using lime or sodium hydroxide as precipitation agent. After pH stabilization and decantation the purified leach liquor and residues are analyzed. The compositions of the purified leach liquors, the final pH and the quantity of lime used in each precipitation are shown in Table 12. Table 12 shows that the majority of iron and aluminum have been precipitated without zinc and manganese.
The compositions of the precipitates are shown in Table 13.
Table 12: Elements concentration in the purified leach liquor (mg/L) pH Ca(OH)2 Mn Pb Al K Ca Fe Mg Sn Zn S Na
(g)
L1 Zn 4~26 4Ϊ7 i~80 704 376 720 O80 5Ϊ0 00 461 2Ϊ09 27Ί i~6 L2 Zn 4.26 856 1 .60 52.9 224 531 0.10 390 0.0 704 2082 10.5 6.3 L3 Zn 4.26 690 1 .60 94.9 103 594 34.9 214 0.0 361 1691 4.30 14.1
Table 13: Elements concentration in the dry precipitates (g/kg)
Moisture
Mn Pb Al K Ca Fe Mg Sn Zn S
%
L1 Zn 93.4 11 .0 1 .41 66.5 64.9 108 2.04 7.78 0.40 7.89 138
L2 Zn 90.5 11 .1 0.43 57.0 28.5 120 3.98 4.26 1.34 8.50 146
L3 Zn 85.9 4.90 0.27 25.7 9.39 138 29.1 1 .63 0.94 3.00 145
EXAMPLE V: ZINC OXIDE PRECIPITATION
Zinc oxide was precipitated from leach liquors after the zinc leaching (see EXAMPLE II) using stoichiometric quantity of NaOH as a precipitation agent (considering the concentration of Zn) at pH between 9.0 and 1 1.0. In the test, 1000 mL of leach liquor L1 Zn with 9339 mg Zn/L is precipitated with 114 mL of 100 g NaOH/L. The results are in Table 14.
Table 14: Initial concentration of elements in leach liquors (mg/L) pH pH
Mn Pb Al K Ca Fe Mg Zn S Na
ini fin
L1 Zn 9.70 10.0 7.20 0.0 460 0.0 39.6 9339 5668 0.0 5.7 10.!
L2 Zn 1 1.5 6.90 24.7 1.4 357 8.3 37.1 1 1536 6845 0.0 5.4 10.0
After the precipitation, 20.0 g of ZnO for L1 Zn and 24.7 g of ZnO for L2 Zn are precipitated. The solid zinc oxide is separated from the solution by a filter. Concentrations of various elements of the finals cake are in Table 15. The concentration of zinc in the cake is 52.9 % for L1 Zn and 50.5 % for L2 Zn.
Table 15: Concentration of elements in g/kg in the final cake
Mn Pb ΑΪ K Ca Fe Mg Zn S Na-
L1 Zn 0.52 0.53 0.48 0.15 8.87 0.02 6.63 529 60.2 48.6 L2 Zn 0.52 0.31 1.13 0.11 9.82 0.39 4.55 505 67.6 55.5
EXAMPLE VI: ZINC SULPHIDE PRECIPITATION
Zinc sulphide was precipitated from purified leach liquors using Na2S as a precipitation agent at pH between about 4.5 and about 5.5. In this test 150 mL of leach liquor L Zn with 855.8 mg Mn/L and 704 mg Zn/L is precipitated with twice the stoichiometry (considering the concentration of Zn) of Na2S. The Mn and Zn concentration after and before the precipitation are in Table 16.
Table 16: Concentration of elements in leach liquor (mg/L)
Mn Pb ΑΪ K Ca Fe Mg Zn S
L Zn
856 1.6 52.9 224 531 0.1 390 704 2082
Before L Zn
855 0.0 50.5 231 476 0.0 343 77.6 2334
After
Table 16 shows that most of zinc is precipitated in sulphides without manganese which will be precipitated as carbonate in another stage.
EXAMPLE VII: MANGANESE CARBONATE PRECIPITATION
Manganese carbonate was precipitated from purified leach liquors using stoichiometry of Na2C03 as a precipitation agent (considering the concentration of Mn) at pH between about 7.5 and 8.5. Before manganese carbonate precipitation, impurities like Al, Fe, Pb and Zn are removed as sulfide from the leach liquor. The pH of the solution is changed to about 4.80 and 5,80 to precipitate aluminum, iron, lead and zinc in a mix tank using sodium hydroxide and sodium sulfide as precipitation agent. After pH stabilization and decantation the purified leach liquor is filtrated, manganese carbonate is then precipitated with Na2C03. Manganese precipitation is very fast at ambient temperature.
In the test, 1000 ml_ of leach liquor L1 Mn with 10125 mg Mn/L is precipitated with 23.4 g of Na2C03. The results are shown in Table 17.
Table 17: Initial concentration in mg/L of elements in leach liquors after
purification pH pH
Mn Pb Al Ca Fe Mg Zn
ini fin
L1
10125 2.1 0.1 23.4 545 1 .1 34.1 0.5 13888 4.8 8.6
Mn
L2
7499 1 .7 0.2 13.8 517 0.6 34.0 0.2 1 1279 5.6 8.6
Mn
After the precipitation, 26.9 g of MnC03 for L1 Mn and 16.4 g of MnC03 for L2 Mn are precipitated. The solid manganous carbonate is separated from the solution by a filter. Concentrations of various elements of the finals cake are in the Table 18. The elements concentration in leach liquors after precipitation of MnC03 are in Table 19, the majority of manganese have been precipitated but calcium, sodium and sulfides are also precipitated.
Table 18: Concentration of elements in g/kg in the final cake Mn Pb ΑΪ K Ca Fe Mg Zn S Na
□
439 0.5 0.1 1 .7 13.2 0.1 0.0 1 .6 17.8 19.3
Mn L2
41 1 0.6 8.9 1 .8 14.5 0.7 0.1 5.3 15.4 15.8
Mn
Table 19: Concentration of elements in leach purified liquor (mg/L)
Mn Pb Al K Ca Fe Mg Zn S Na
L1
5.6 1 .1 0.0 21 .6 58.2 0.0 29.0 0.0 1 1727 13518
Mn
6.2 0.7 0.1 14.4 36.8 0.6 25.4 0.0 1 1722 10259
L2
Mn
After the precipitation, it is preferred to wash the manganous carbonate precipitates to eliminate the sodium and sulfides. In the test, 20 g of manganous carbonate was rinsed with 200 ml_ of water (10% pulp density by weight), the results are shown in Table 20. The concentration of Mn in the cake is 44.6 % for L1 and 43.0 % for L2.
Table 20: Concentration of elements in g/kg in the final cake after
rinsing
Mn Pb Al K Ca Fe Mg Zn S Na
L1
446 0,3 0,0 1.9 20.9 0,0 0,3 0.0 2.3 7,6
Mn
L2
430 0,2 0.0 0.7 28.9 0,0 0,7 0.0 1.8 6.4
Mn
EXEMPLE VIII: ZN ELECTROPLATING
Zn metal was electrodeposited from leach liquor using an electroplating cell. An electrical power source is connected to four cathodes and four anodes. The cathodes are made of stainless steel and the anodes are made of Ti02. The area of each cathode is 220 cm2. The solution is fed into electrolytic tank of 2 liters where zinc metal is deposited on the stainless steel cathodes during electrolysis. At the Ti02 anodes an oxidation reaction occurs, generating oxygen gas, protons and giving electrons. The current is fixed of 15 A and every 60 minutes the cathodes full of Zn metal are replaced by news cathodes.
Reduction at cathode: Zn2+ + 2e" Zn
Oxidation at anode: H20 2H+ + ½02 + 2e"
Zinc metal was electrolysed from the mixture of leach liquors L1 and L2 after the zinc leaching (see EXAMPLE II). The results are shown in Table 21 and Figure 1 shows the Zn concentration during the electrolysis.
Table 21 : Concentration of elements in leach liquors (mg/L)
Mn Pb Al K Ca Fe Mg Zn S Na
L1 and L2 Zn
8.2 7.7 17.3 0.0 240 14.6 33.5 10966 5713 0.0
Before
L1 and L2 Zn
2.9 0.0 20.8 0.0 186 12.3 26.7 1530 5714 0.0
After
With four change of cathode, about 86.0% of Zn is electrolyzed and the final concentration of Zn reaches 1530 mg/L.
While the techniques described herein apply to steel plant dust as well as various streams and materials derived therefrom, they can also be used for treating other matrices to extract valuable materials. For example, various waste or byproduct streams from the metals processing industry may be processed to extract compounds such as Zn, Mn and/or Pb, in particular when the concentration of such compounds is economically significant. Waste streams can include electric arc furnace (EAF) dust, various pyrometallurgical residues, and materials derived therefrom. Waste streams from the metals processing industry or other matrices can also be tested to determine the composition and extractability of desired components, prior to subjecting the waste streams to an implementation of the techniques described herein. Various other matrices, such as mining residues, can also be treated with techniques described herein.
Claims
1. A process for treating a matrix comprising zinc (Zn), manganese (Mn) and lead (Pb), comprising: leaching the matrix to generate a leaching solution enriched in Zn and a fraction enriched in Mn and Pb; subjecting the leaching solution to Zn-precipitation to produce Zn-based precipitates and a Zn-depleted fraction comprising Mn compounds; leaching the fraction enriched in Mn and Pb to produce a Mn-enriched leaching solution and a Pb-enriched fraction; and subjecting the Zn-depleted fraction and/or the Mn-enriched leaching solution to Mn-precipitation to produce Mn-based precipitates.
2. The process of claim 1 , wherein the leaching of the matrix is performed using an aqueous acidic solution.
3. The process of claim 2, wherein the aqueous acidic solution comprises H2S04.
4. The process of claim 2 or 3, wherein the leaching of the matrix comprises multiple leaching runs with the aqueous acidic solution and multiple rinsing runs with an aqueous rinsing solution.
5. The process of claim 4, wherein the leaching of the matrix comprises at least three leaching runs and/or at least three rinsing runs.
6. The process of any one of claims 1 to 5, further comprising, prior to the Zn- precipitation, removing additional metal components from the leaching solution enriched in Zn.
7. The process of claim 6, wherein the additional metal components include aluminum (Al).
8. The process of claim 6 or 7, wherein the additional metal components include iron (Fe).
9. The process of any one of claims 6 to 8, wherein the removing of the additional metal components comprises precipitating the additional metal components.
10. The process of claim 9, wherein precipitating of the additional metal components comprises adding a hydroxide of a divalent cation.
1 1. The process of claim 10, wherein the hydroxide of a divalent cation comprises Ca(OH)2 or NaOH.
12. The process of any one of claims 1 to 1 1 , wherein the Zn-precipitation comprises adding NaOH or Na2S to produce ZnO or ZnS as the Zn-based precipitates.
13. The process according to any one of claims 1 to 12, further comprising, prior to the Mn-precipitation, removing additional metallic impurities from the Zn-depleted fraction and/or the Mn-enriched leaching solution.
14. ,The process of claim 13, wherein the additional metallic impurities comprise aluminum, iron, lead, zinc or a combination thereof.
15. The process of claim 13 or 14, wherein removing the additional metallic impurities prior to the Mn-precipitation comprises adding NaOH and Na2S to the Zn-depleted fraction and/or the Mn-enriched leaching solution.
16. The process of any one of claims 1 to 15, wherein the Mn-precipitation comprises adding a carbonate compound to produce Mn-carbonates as the Mn-based precipitates.
17. The process of claim 16, wherein the carbonate compound comprises Na2C03 to produce MnC03.
18. The process of any one of claims 1 to 17, further comprising, prior to the leaching of the matrix, a step of determining a composition of the matrix.
19. The process of claim 18, wherein determining the composition of the matrix comprises determining the concentrations of Zn and Mn.
20. The process of claim 18 or 19, wherein determining the composition of the matrix comprises determining the concentration of Pb.
21. The process of any one of claims 18 to 20, wherein determining the composition of the matrix includes determining the concentrations of Al and/or Fe.
22. The process of any one of claims 1 to 21 , further comprising treating the Pb-enriched fraction to recover Pb therefrom.
23. The process of any one of claims 1 to 22, further comprising: subjecting the Zn-depleted fraction to Mn-precipitation in a first reactor to produce a first stream of Mn-based precipitates; and subjecting the the Mn-enriched leaching solution to Mn-precipitation in a second reactor to produce a second stream of Mn-based precipitates.
24. The process of claim 23, further comprising: producing a Mn-depleted stream in the second reactor; and subjecting at least a portion of the Mn-depleted stream to further Mn-precipitation.
25. The process of claim 24, wherein the further Mn-precipitation of the Mn-depleted stream is conducted in the first reactor.
26. The process of any one of claims 23 to 25, further comprising recycling a portion of the Mn-depleted stream from the second reactor for addition to the matrix before or during the leaching thereof.
27. The process of any one of claims 23 to 26, further comprising recovering a metals- depleted liquid from the Mn-precipitation in the first reactor and adding at least a portion of the metals-depleted liquid to the matrix before or during the leaching thereof.
28. The process of any one of claims 1 to 27, wherein the leaching of the matrix is performed in atmospheric pressure conditions.
29. The process of any one of claims 1 to 27, wherein one or more of the leaching steps are performed in a counter-current reactor.
30. The process of any one of claims 1 to 28, wherein one or more of the leaching steps are performed in an agitated tank reactor.
31. The process of any one of claims 1 to 28, wherein one or more of the leaching steps are performed in a batch reactor or a continuous.
32. The process of any one of claims 1 to 31 , wherein the matrix comprises a powder or dust.
33. The process of any one of claims 1 to 32, wherein the matrix comprises a by-product or waste stream derived from metals processing.
34. The process of any one of claims 1 to 33, wherein the matrix comprises steel plant dust.
35. The process of any one of claims 1 to 34, including one or more features of claims 33 to 53.
36. A process for treating a matrix comprising zinc (Zn), manganese (Mn) and lead (Pb), comprising selective and sequential leaching of Zn, Mn and Pb in order to obtain a Mn concentrate, a Zn concentrate, and a Pb-enriched residue.
37. The process of claim 36, including one or more features of claims 1 to 35 and 38 to 56.
38. A method for the hydrometallurgical treatment of a matrix comprising manganese (Mn), zinc (Zn), lead (Pb), iron (Fe), aluminum (Al), and at least one of calcium, chlorides, sodium and potassium as impurities, the method comprising: a first leaching step comprising: leaching the matrix at atmospheric pressure in a reactor using a first leaching solution to product a first leach liquor and a first leach residue, the first leaching solution comprising a sulfuric acid solution; and separating the first leach liquor from the first leach residue after the first leaching;
a second leaching step comprising: leaching the first leach residue at atmospheric pressure using a second leaching solution to form a second leach liquor and a second leach residue, the second leaching solution comprising sulfuric acid and a reducing agent; and separating the second leach liquor from the second leach residue; a Mn-precipitation step wherein Mn compounds are precipitated from the first and second leach liquors, including contacting the liquors or streams derived therefrom, with a carbonate at a pH to precipitate Mn carbonate compounds.
39. The method of claim 38, further comprising selecting a quantity of the carbonate to adjust the pH of the first and second leach liquor sufficiently to precipitate out Mn carbonate compounds.
40. The method of claim 38 or 39, wherein the pH of the second leaching is between about 3 and about 5.
41. The method of any one of claims 38 to 40, wherein the reducing agent used in the second leaching step comprises Na2S205.
42. The method of any one of claims 38 to 41 , further comprising subjecting the first liquor to Zn-precipitation prior to Mn-precipitation.
43. The method of claim 42, wherein the Zn-precipitation includes electroplating Zn as Zn metal.
44. The method of claim 42, wherein the Zn-precipitation includes forming Zn sulfide and/or Zn oxide precipitates that are precipitated out of the solution.
45. The method of claim 44, further comprising separating the Zn sulfide and/or Zn oxide precipitates from the solution to form a Zn sulfide and/or Zn oxide containing residue, and a Zn-depleted first liquor that is subjected to the Mn-precipitation step.
46. The method of claim 44 or 45, wherein the Zn-precipitation comprises:
adding alkali or sulfide to the first liquor to precipitate Zn compounds; and then preforming a solid-liquid separation to produce a Zn sulfide and/or oxide cake.
47. The method of any one of claims 44 to 46, further comprising: subjecting the first liquor to a Fe, silicate and Al precipitation step to form Fe, silicate and Al hydroxide compounds and a depleted liquor; separating the Fe, silicate and Al hydroxide compounds from the depleted liquor; and subjecting the depleted liquor to the Zn-precipitation.
48. The method of the claim 47, wherein the Fe, silicate and Al precipitation step comprises: adding alkali to increase the pH between about 4 and about 5; and then preforming a solid-liquid separation to produce a mixed Fe, silicate and Al hydroxide cake.
49. The method of any one of claims 38 to 48, wherein the Mn-precipitation step comprises: adding the carbonate to the first and second liquors to increase the pH to between about 7.5 and about 8.5; and then performing a solid-liquid separation to produce a Mn-carbonate cake.
50. The method of any one of claims 38 to 49, wherein the second leach residue contains between 10 wt% and 30 wt% of Pb.
51. The method of any one of claims 38 to 50, wherein a leaching period of mixing the sulfuric acid solution for the first and/or second leaching steps is between about 10 minutes and about 24 h.
52. The method of any one of claims 38 to 51 , wherein the temperature of the sulfuric acid solution is between about 20°C and about 40°C.
53. The method of any one of claims 38 to 52, further comprising one or more features of claims 1 to 34.
54. A system for treating a matrix comprising zinc (Zn) and manganese (Mn), comprising: a leaching unit for leaching the matrix to generate a Zn-enriched leaching solution, and a fraction enriched in Mn and Pb; a Zn-precipitation unit for subjecting the Zn-enriched leaching solution to Zn- precipitation to produce Zn-based precipitates and a Zn-depleted fraction comprising Mn compounds; and a leaching unit for leaching the fraction enriched in Mn and Pb to produce a Mn- enriched leaching solution and a Pb-enriched fraction; and a Mn-precipitation system for subjecting the Zn-depleted fraction and/or the Mn- enriched leaching solution, or streams derived therefrom, to Mn-precipitation to produce Mn-based precipitates.
55. The system of claim 54, wherein the Mn-precipitation system comprises: a first Mn-precipitation unit for subjecting the Zn-depleted fraction to Mn- precipitation; and a second Mn-precipitation unit for subjecting the Mn-enriched leaching solution to Mn-precipitation.
56. The system of claim 54 or 55, further comprising an additional precipitation unit for receiving the Zn-enriched leaching solution and removing at least one of aluminum (Al), iron (Fe) and silicate therefrom, to produce a liquor for supplying to the Zn- precipitation unit.
57. A system for the hydrometallurgical treatment of a matrix comprising manganese (Mn), zinc (Zn), lead (Pb), iron (Fe), aluminum (Al), and at least one of calcium, chlorides, sodium and potassium as impurities, the system comprising:
a first leaching unit for leaching the matrix at atmospheric pressure in a reactor using a first leaching solution to product a first leach liquor and a first leach residue, the first leaching solution comprising a sulfuric acid solution; a first separation unit for separating the first leach liquor from the first leach residue after the first leaching; a second leaching unit for leaching the first leach residue at atmospheric pressure using a second leaching solution to form a second leach liquor and a second leach residue, the second leaching solution comprising sulfuric acid and a reducing agent; a second separation unit for separating the second leach liquor from the second leach residue; and a Mn-precipitation system for precipitating Mn compounds from the first and second leach liquors, or streams derived therefrom, by contacting with a carbonate at a pH to precipitate Mn carbonate compounds.
58. The system of claim 57, wherein the Mn-precipitation system comprises: a first Mn-precipitation unit for subjecting the first leach liquor to Mn-precipitation; and a second Mn-precipitation unit for subjecting the second leach liquor to Mn- precipitation.
59. The system of claim 58, further comprising: an additional precipitation unit for receiving the first leach liquor and removing at least one of aluminum (Al), iron (Fe) and silicate therefrom, to produce a metals depleted liquor comprising Zn and Mn; a Zn-precipitation unit for receiving the metals depleted liquor and producing Zn precipitates and a Zn-depleted fraction, wherein the first precipitation unit is configured to receive the Zn-depleted fraction to be subjected to the Mn- precipitation.
60. The system of any one of claims 54 to 59, further comprising one or more operational or structural features of claims 1 to 53 and/or described in the description or drawings.
Applications Claiming Priority (2)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA2854778A CA2854778A1 (en) | 2014-06-18 | 2014-06-18 | Recovery of zinc and manganese from pyrometalurgy sludge or residues |
| PCT/CA2015/050547 WO2015192234A1 (en) | 2014-06-18 | 2015-06-12 | Recovery of zinc and manganese from pyrometallurgy sludge or residues |
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| EP3172348A1 true EP3172348A1 (en) | 2017-05-31 |
| EP3172348A4 EP3172348A4 (en) | 2018-05-23 |
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| EP15809510.9A Withdrawn EP3172348A4 (en) | 2014-06-18 | 2015-06-12 | Recovery of zinc and manganese from pyrometallurgy sludge or residues |
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| FI128281B (en) * | 2016-12-15 | 2020-02-28 | Teknologian Tutkimuskeskus Vtt Oy | Processing of Industrial Metal-Containing Waste Materials |
| CN110436721A (en) * | 2019-08-12 | 2019-11-12 | 江苏荣信环保科技有限公司 | A kind of metal surface pickling sludge and spent acid comprehensive processing technique |
| CN112111654B (en) * | 2020-10-09 | 2021-10-22 | 昆明理工大学 | Method for separating lead and manganese from electrolytic manganese anode slime |
| CN113061736B (en) * | 2021-03-30 | 2022-03-22 | 攀钢集团攀枝花钢铁研究院有限公司 | Method for separating potassium, lead and iron from sintering machine head ash |
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| GB1290672A (en) * | 1969-03-07 | 1972-09-27 | ||
| GB1509537A (en) * | 1974-09-13 | 1978-05-04 | Cominco Ltd | Treatment of zinc plant residues |
| US5534234A (en) * | 1994-11-14 | 1996-07-09 | Reddin; Lorin D. | Recovery of manganese from leach solutions |
| TR200707718T1 (en) * | 2005-05-10 | 2008-04-21 | Puwada George | Process for treating powders and residues of electric and other furnaces containing zinc oxide and zinc ferrite. |
| WO2009157620A1 (en) * | 2008-06-25 | 2009-12-30 | Dongbu Fine Chemicals Co., Ltd. | Method of recovering a compound comprising manganese from dust of electronic furnace |
| KR101011260B1 (en) * | 2009-10-27 | 2011-01-26 | 주식회사 에코닉스 | Method for making cmd from waste batteries containing manganese and zinc |
| US8460631B2 (en) * | 2010-08-18 | 2013-06-11 | American Manganese Inc. | Processing of manganous sulphate/dithionate liquors derived from manganese resource material |
| CN102586599B (en) * | 2012-03-07 | 2013-07-31 | 株洲冶炼集团股份有限公司 | Method for recovering valued metals from zinc anode sludge |
-
2014
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