JP2005060813A - Method for refining copper raw material containing copper sulfide mineral - Google Patents
Method for refining copper raw material containing copper sulfide mineral Download PDFInfo
- Publication number
- JP2005060813A JP2005060813A JP2003315124A JP2003315124A JP2005060813A JP 2005060813 A JP2005060813 A JP 2005060813A JP 2003315124 A JP2003315124 A JP 2003315124A JP 2003315124 A JP2003315124 A JP 2003315124A JP 2005060813 A JP2005060813 A JP 2005060813A
- Authority
- JP
- Japan
- Prior art keywords
- copper
- leaching
- raw material
- refining
- iron
- 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
- 239000010949 copper Substances 0.000 title claims abstract description 325
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 title claims abstract description 309
- 229910052802 copper Inorganic materials 0.000 title claims abstract description 306
- 238000000034 method Methods 0.000 title claims abstract description 217
- 238000007670 refining Methods 0.000 title claims abstract description 93
- OMZSGWSJDCOLKM-UHFFFAOYSA-N copper(II) sulfide Chemical compound [S-2].[Cu+2] OMZSGWSJDCOLKM-UHFFFAOYSA-N 0.000 title claims abstract description 90
- 229910052569 sulfide mineral Inorganic materials 0.000 title claims abstract description 90
- 239000002994 raw material Substances 0.000 title claims abstract description 89
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims abstract description 238
- 238000002386 leaching Methods 0.000 claims abstract description 179
- 239000007788 liquid Substances 0.000 claims abstract description 179
- 229910052742 iron Inorganic materials 0.000 claims abstract description 124
- 230000009467 reduction Effects 0.000 claims abstract description 90
- 230000008569 process Effects 0.000 claims abstract description 69
- JPVYNHNXODAKFH-UHFFFAOYSA-N Cu2+ Chemical compound [Cu+2] JPVYNHNXODAKFH-UHFFFAOYSA-N 0.000 claims abstract description 67
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 claims abstract description 64
- 229910001431 copper ion Inorganic materials 0.000 claims abstract description 64
- 239000000460 chlorine Substances 0.000 claims abstract description 62
- 229910052801 chlorine Inorganic materials 0.000 claims abstract description 62
- 238000005363 electrowinning Methods 0.000 claims abstract description 62
- ZAMOUSCENKQFHK-UHFFFAOYSA-N Chlorine atom Chemical compound [Cl] ZAMOUSCENKQFHK-UHFFFAOYSA-N 0.000 claims abstract description 61
- 238000000638 solvent extraction Methods 0.000 claims abstract description 61
- 229910052717 sulfur Inorganic materials 0.000 claims abstract description 43
- 239000011593 sulfur Substances 0.000 claims abstract description 43
- 229910052951 chalcopyrite Inorganic materials 0.000 claims abstract description 31
- DVRDHUBQLOKMHZ-UHFFFAOYSA-N chalcopyrite Chemical compound [S-2].[S-2].[Fe+2].[Cu+2] DVRDHUBQLOKMHZ-UHFFFAOYSA-N 0.000 claims abstract description 31
- 239000002699 waste material Substances 0.000 claims abstract description 28
- 238000000746 purification Methods 0.000 claims abstract description 25
- 229910052751 metal Inorganic materials 0.000 claims abstract description 24
- 239000002184 metal Substances 0.000 claims abstract description 24
- 238000005868 electrolysis reaction Methods 0.000 claims abstract description 22
- 239000003638 chemical reducing agent Substances 0.000 claims abstract description 11
- 239000000243 solution Substances 0.000 claims description 116
- 239000000047 product Substances 0.000 claims description 99
- 238000000605 extraction Methods 0.000 claims description 76
- 239000012141 concentrate Substances 0.000 claims description 51
- VEXZGXHMUGYJMC-UHFFFAOYSA-M Chloride anion Chemical compound [Cl-] VEXZGXHMUGYJMC-UHFFFAOYSA-M 0.000 claims description 42
- 229910021607 Silver chloride Inorganic materials 0.000 claims description 38
- HKZLPVFGJNLROG-UHFFFAOYSA-M silver monochloride Chemical compound [Cl-].[Ag+] HKZLPVFGJNLROG-UHFFFAOYSA-M 0.000 claims description 38
- 150000002500 ions Chemical class 0.000 claims description 33
- 239000002245 particle Substances 0.000 claims description 32
- 239000007864 aqueous solution Substances 0.000 claims description 28
- -1 iron ions Chemical class 0.000 claims description 24
- KZBUYRJDOAKODT-UHFFFAOYSA-N Chlorine Chemical compound ClCl KZBUYRJDOAKODT-UHFFFAOYSA-N 0.000 claims description 20
- 239000002002 slurry Substances 0.000 claims description 18
- 229910052500 inorganic mineral Inorganic materials 0.000 claims description 17
- 239000011707 mineral Substances 0.000 claims description 17
- 239000003960 organic solvent Substances 0.000 claims description 15
- FBAFATDZDUQKNH-UHFFFAOYSA-M iron chloride Chemical compound [Cl-].[Fe] FBAFATDZDUQKNH-UHFFFAOYSA-M 0.000 claims description 13
- 150000002739 metals Chemical class 0.000 claims description 13
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims description 12
- CWYNVVGOOAEACU-UHFFFAOYSA-N Fe2+ Chemical compound [Fe+2] CWYNVVGOOAEACU-UHFFFAOYSA-N 0.000 claims description 11
- 229910001448 ferrous ion Inorganic materials 0.000 claims description 11
- 239000004744 fabric Substances 0.000 claims description 9
- 229910052709 silver Inorganic materials 0.000 claims description 9
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 claims description 8
- 239000004332 silver Substances 0.000 claims description 8
- 238000009825 accumulation Methods 0.000 claims description 7
- 230000002378 acidificating effect Effects 0.000 claims description 6
- 238000004821 distillation Methods 0.000 claims description 6
- 230000007935 neutral effect Effects 0.000 claims description 6
- 238000006386 neutralization reaction Methods 0.000 claims description 6
- 230000033116 oxidation-reduction process Effects 0.000 claims description 6
- 230000035699 permeability Effects 0.000 claims description 5
- NGNBDVOYPDDBFK-UHFFFAOYSA-N 2-[2,4-di(pentan-2-yl)phenoxy]acetyl chloride Chemical compound CCCC(C)C1=CC=C(OCC(Cl)=O)C(C(C)CCC)=C1 NGNBDVOYPDDBFK-UHFFFAOYSA-N 0.000 claims description 4
- 239000011260 aqueous acid Substances 0.000 claims description 4
- 238000010438 heat treatment Methods 0.000 claims description 4
- 239000010970 precious metal Substances 0.000 claims description 4
- 238000005987 sulfurization reaction Methods 0.000 claims description 4
- 238000007664 blowing Methods 0.000 claims description 3
- 238000005486 sulfidation Methods 0.000 claims description 3
- 238000007743 anodising Methods 0.000 claims description 2
- 229910001779 copper mineral Inorganic materials 0.000 claims description 2
- 238000000151 deposition Methods 0.000 claims description 2
- 239000000463 material Substances 0.000 claims description 2
- 238000001556 precipitation Methods 0.000 claims 1
- 238000007254 oxidation reaction Methods 0.000 abstract description 27
- 230000003647 oxidation Effects 0.000 abstract description 25
- 238000011084 recovery Methods 0.000 abstract description 8
- 239000007787 solid Substances 0.000 abstract description 6
- VMQMZMRVKUZKQL-UHFFFAOYSA-N Cu+ Chemical compound [Cu+] VMQMZMRVKUZKQL-UHFFFAOYSA-N 0.000 abstract description 4
- 238000004519 manufacturing process Methods 0.000 abstract description 2
- 238000006722 reduction reaction Methods 0.000 description 86
- STCOOQWBFONSKY-UHFFFAOYSA-N tributyl phosphate Chemical compound CCCCOP(=O)(OCCCC)OCCCC STCOOQWBFONSKY-UHFFFAOYSA-N 0.000 description 24
- 238000006243 chemical reaction Methods 0.000 description 20
- 235000010755 mineral Nutrition 0.000 description 15
- 239000000203 mixture Substances 0.000 description 13
- 239000012535 impurity Substances 0.000 description 12
- 230000001590 oxidative effect Effects 0.000 description 12
- 238000003723 Smelting Methods 0.000 description 11
- ORTQZVOHEJQUHG-UHFFFAOYSA-L copper(II) chloride Chemical compound Cl[Cu]Cl ORTQZVOHEJQUHG-UHFFFAOYSA-L 0.000 description 11
- FAPWRFPIFSIZLT-UHFFFAOYSA-M Sodium chloride Chemical compound [Na+].[Cl-] FAPWRFPIFSIZLT-UHFFFAOYSA-M 0.000 description 10
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 description 10
- 229910021578 Iron(III) chloride Inorganic materials 0.000 description 9
- 239000011324 bead Substances 0.000 description 9
- RBTARNINKXHZNM-UHFFFAOYSA-K iron trichloride Chemical compound Cl[Fe](Cl)Cl RBTARNINKXHZNM-UHFFFAOYSA-K 0.000 description 9
- 238000010298 pulverizing process Methods 0.000 description 9
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 8
- 238000004458 analytical method Methods 0.000 description 8
- 238000002156 mixing Methods 0.000 description 8
- IJGRMHOSHXDMSA-UHFFFAOYSA-N Atomic nitrogen Chemical compound N#N IJGRMHOSHXDMSA-UHFFFAOYSA-N 0.000 description 7
- VTLYFUHAOXGGBS-UHFFFAOYSA-N Fe3+ Chemical compound [Fe+3] VTLYFUHAOXGGBS-UHFFFAOYSA-N 0.000 description 7
- 229910001447 ferric ion Inorganic materials 0.000 description 7
- 229910000510 noble metal Inorganic materials 0.000 description 7
- PXHVJJICTQNCMI-UHFFFAOYSA-N Nickel Chemical compound [Ni] PXHVJJICTQNCMI-UHFFFAOYSA-N 0.000 description 6
- 239000012071 phase Substances 0.000 description 6
- 238000004070 electrodeposition Methods 0.000 description 5
- 238000000227 grinding Methods 0.000 description 5
- 229910021645 metal ion Inorganic materials 0.000 description 5
- 239000011780 sodium chloride Substances 0.000 description 5
- 239000000126 substance Substances 0.000 description 5
- 229910021591 Copper(I) chloride Inorganic materials 0.000 description 4
- 229910052787 antimony Inorganic materials 0.000 description 4
- WATWJIUSRGPENY-UHFFFAOYSA-N antimony atom Chemical compound [Sb] WATWJIUSRGPENY-UHFFFAOYSA-N 0.000 description 4
- 229910052785 arsenic Inorganic materials 0.000 description 4
- RQNWIZPPADIBDY-UHFFFAOYSA-N arsenic atom Chemical compound [As] RQNWIZPPADIBDY-UHFFFAOYSA-N 0.000 description 4
- OXBLHERUFWYNTN-UHFFFAOYSA-M copper(I) chloride Chemical compound [Cu]Cl OXBLHERUFWYNTN-UHFFFAOYSA-M 0.000 description 4
- 229940045803 cuprous chloride Drugs 0.000 description 4
- 239000007789 gas Substances 0.000 description 4
- 238000000926 separation method Methods 0.000 description 4
- RTAQQCXQSZGOHL-UHFFFAOYSA-N Titanium Chemical compound [Ti] RTAQQCXQSZGOHL-UHFFFAOYSA-N 0.000 description 3
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 description 3
- 239000008346 aqueous phase Substances 0.000 description 3
- 229910052947 chalcocite Inorganic materials 0.000 description 3
- 229960003280 cupric chloride Drugs 0.000 description 3
- 239000012527 feed solution Substances 0.000 description 3
- 238000005259 measurement Methods 0.000 description 3
- 229910052759 nickel Inorganic materials 0.000 description 3
- 229910052757 nitrogen Inorganic materials 0.000 description 3
- 238000011946 reduction process Methods 0.000 description 3
- 239000010936 titanium Substances 0.000 description 3
- 229910052725 zinc Inorganic materials 0.000 description 3
- 239000011701 zinc Substances 0.000 description 3
- QPLDLSVMHZLSFG-UHFFFAOYSA-N Copper oxide Chemical compound [Cu]=O QPLDLSVMHZLSFG-UHFFFAOYSA-N 0.000 description 2
- 239000005751 Copper oxide Substances 0.000 description 2
- BUGBHKTXTAQXES-UHFFFAOYSA-N Selenium Chemical compound [Se] BUGBHKTXTAQXES-UHFFFAOYSA-N 0.000 description 2
- QAOWNCQODCNURD-UHFFFAOYSA-L Sulfate Chemical compound [O-]S([O-])(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-L 0.000 description 2
- 229910052797 bismuth Inorganic materials 0.000 description 2
- JCXGWMGPZLAOME-UHFFFAOYSA-N bismuth atom Chemical compound [Bi] JCXGWMGPZLAOME-UHFFFAOYSA-N 0.000 description 2
- 238000009835 boiling Methods 0.000 description 2
- 229910052793 cadmium Inorganic materials 0.000 description 2
- BDOSMKKIYDKNTQ-UHFFFAOYSA-N cadmium atom Chemical compound [Cd] BDOSMKKIYDKNTQ-UHFFFAOYSA-N 0.000 description 2
- 230000008859 change Effects 0.000 description 2
- 238000001816 cooling Methods 0.000 description 2
- BUGICWZUDIWQRQ-UHFFFAOYSA-N copper iron sulfane Chemical compound S.[Fe].[Cu] BUGICWZUDIWQRQ-UHFFFAOYSA-N 0.000 description 2
- 229910000431 copper oxide Inorganic materials 0.000 description 2
- 239000003085 diluting agent Substances 0.000 description 2
- 238000009826 distribution Methods 0.000 description 2
- 239000008151 electrolyte solution Substances 0.000 description 2
- 238000005188 flotation Methods 0.000 description 2
- 229910021478 group 5 element Inorganic materials 0.000 description 2
- 229910021476 group 6 element Inorganic materials 0.000 description 2
- 239000003350 kerosene Substances 0.000 description 2
- 239000012528 membrane Substances 0.000 description 2
- 239000007800 oxidant agent Substances 0.000 description 2
- 230000009257 reactivity Effects 0.000 description 2
- 229910052711 selenium Inorganic materials 0.000 description 2
- 239000011669 selenium Substances 0.000 description 2
- SQGYOTSLMSWVJD-UHFFFAOYSA-N silver(1+) nitrate Chemical compound [Ag+].[O-]N(=O)=O SQGYOTSLMSWVJD-UHFFFAOYSA-N 0.000 description 2
- 239000007784 solid electrolyte Substances 0.000 description 2
- 239000002904 solvent Substances 0.000 description 2
- 230000001629 suppression Effects 0.000 description 2
- 229910052714 tellurium Inorganic materials 0.000 description 2
- PORWMNRCUJJQNO-UHFFFAOYSA-N tellurium atom Chemical compound [Te] PORWMNRCUJJQNO-UHFFFAOYSA-N 0.000 description 2
- 229910052719 titanium Inorganic materials 0.000 description 2
- 229910001200 Ferrotitanium Inorganic materials 0.000 description 1
- MBMLMWLHJBBADN-UHFFFAOYSA-N Ferrous sulfide Chemical compound [Fe]=S MBMLMWLHJBBADN-UHFFFAOYSA-N 0.000 description 1
- 229910000831 Steel Inorganic materials 0.000 description 1
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 description 1
- IDCBOTIENDVCBQ-UHFFFAOYSA-N TEPP Chemical compound CCOP(=O)(OCC)OP(=O)(OCC)OCC IDCBOTIENDVCBQ-UHFFFAOYSA-N 0.000 description 1
- ATJFFYVFTNAWJD-UHFFFAOYSA-N Tin Chemical compound [Sn] ATJFFYVFTNAWJD-UHFFFAOYSA-N 0.000 description 1
- MKOYQDCOZXHZSO-UHFFFAOYSA-N [Cu].[Cu].[Cu].[As] Chemical compound [Cu].[Cu].[Cu].[As] MKOYQDCOZXHZSO-UHFFFAOYSA-N 0.000 description 1
- 239000002253 acid Substances 0.000 description 1
- KGHMFMDJVUVBRY-UHFFFAOYSA-N antimony copper Chemical compound [Cu].[Sb] KGHMFMDJVUVBRY-UHFFFAOYSA-N 0.000 description 1
- 239000010953 base metal Substances 0.000 description 1
- 229940006460 bromide ion Drugs 0.000 description 1
- CODNYICXDISAEA-UHFFFAOYSA-N bromine monochloride Chemical compound BrCl CODNYICXDISAEA-UHFFFAOYSA-N 0.000 description 1
- AXCZMVOFGPJBDE-UHFFFAOYSA-L calcium dihydroxide Chemical compound [OH-].[OH-].[Ca+2] AXCZMVOFGPJBDE-UHFFFAOYSA-L 0.000 description 1
- 239000000920 calcium hydroxide Substances 0.000 description 1
- 235000011116 calcium hydroxide Nutrition 0.000 description 1
- 229910001861 calcium hydroxide Inorganic materials 0.000 description 1
- 238000006388 chemical passivation reaction Methods 0.000 description 1
- 239000003153 chemical reaction reagent Substances 0.000 description 1
- 238000005660 chlorination reaction Methods 0.000 description 1
- 150000001805 chlorine compounds Chemical class 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
- 238000006477 desulfuration reaction Methods 0.000 description 1
- 230000023556 desulfurization Effects 0.000 description 1
- 230000006866 deterioration Effects 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000007865 diluting Methods 0.000 description 1
- 229910001873 dinitrogen Inorganic materials 0.000 description 1
- 238000007599 discharging Methods 0.000 description 1
- 230000000694 effects Effects 0.000 description 1
- 230000005611 electricity Effects 0.000 description 1
- 239000003792 electrolyte Substances 0.000 description 1
- 229960002089 ferrous chloride Drugs 0.000 description 1
- 239000011521 glass Substances 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
- 150000004820 halides Chemical class 0.000 description 1
- 238000009854 hydrometallurgy Methods 0.000 description 1
- 238000009434 installation Methods 0.000 description 1
- NMCUIPGRVMDVDB-UHFFFAOYSA-L iron dichloride Chemical compound Cl[Fe]Cl NMCUIPGRVMDVDB-UHFFFAOYSA-L 0.000 description 1
- 239000012633 leachable Substances 0.000 description 1
- 239000011133 lead Substances 0.000 description 1
- 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 1
- 229910052592 oxide mineral Inorganic materials 0.000 description 1
- 238000003918 potentiometric titration Methods 0.000 description 1
- 239000002244 precipitate Substances 0.000 description 1
- 238000003672 processing method Methods 0.000 description 1
- 229910052683 pyrite Inorganic materials 0.000 description 1
- NIFIFKQPDTWWGU-UHFFFAOYSA-N pyrite Chemical compound [Fe+2].[S-][S-] NIFIFKQPDTWWGU-UHFFFAOYSA-N 0.000 description 1
- 239000011028 pyrite Substances 0.000 description 1
- 229910052952 pyrrhotite Inorganic materials 0.000 description 1
- 239000010453 quartz Substances 0.000 description 1
- 150000003839 salts Chemical class 0.000 description 1
- 238000005070 sampling Methods 0.000 description 1
- 238000004062 sedimentation Methods 0.000 description 1
- 229910052604 silicate mineral Inorganic materials 0.000 description 1
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N silicon dioxide Inorganic materials O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 description 1
- 239000010944 silver (metal) Substances 0.000 description 1
- 229910001961 silver nitrate Inorganic materials 0.000 description 1
- 239000002893 slag Substances 0.000 description 1
- HYHCSLBZRBJJCH-UHFFFAOYSA-M sodium hydrosulfide Chemical compound [Na+].[SH-] HYHCSLBZRBJJCH-UHFFFAOYSA-M 0.000 description 1
- 229910001415 sodium ion Inorganic materials 0.000 description 1
- 229910001220 stainless steel Inorganic materials 0.000 description 1
- 239000010935 stainless steel Substances 0.000 description 1
- 239000010959 steel Substances 0.000 description 1
- 238000003860 storage Methods 0.000 description 1
- 150000004763 sulfides Chemical class 0.000 description 1
- 229910052718 tin Inorganic materials 0.000 description 1
- 230000005068 transpiration Effects 0.000 description 1
- 238000001238 wet grinding Methods 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
- C22B15/00—Obtaining copper
- C22B15/0063—Hydrometallurgy
- C22B15/0065—Leaching or slurrying
- C22B15/0067—Leaching or slurrying with acids or salts thereof
- C22B15/0069—Leaching or slurrying with acids or salts thereof containing halogen
-
- 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
- C22B15/0084—Treating solutions
- C22B15/0089—Treating solutions by chemical methods
-
- 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
- C22B15/0095—Process control or regulation methods
- C22B15/0097—Sulfur release abatement
-
- 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
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- Engineering & Computer Science (AREA)
- Chemical & Material Sciences (AREA)
- Manufacturing & Machinery (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Metallurgy (AREA)
- Organic Chemistry (AREA)
- Chemical Kinetics & Catalysis (AREA)
- General Chemical & Material Sciences (AREA)
- Automation & Control Theory (AREA)
- Manufacture And Refinement Of Metals (AREA)
- Electrolytic Production Of Metals (AREA)
Abstract
Description
本発明は、硫化銅鉱物を含む銅原料の精錬方法に関し、さらに詳しくは、黄銅鉱を始めとする硫化銅鉱物を含む銅原料の湿式精錬法において、硫黄の酸化を抑制しながら高浸出率で銅を浸出して一価銅電解で回収し、また随伴する有価金属も回収して、浸出残渣などの廃棄物を可能な限り減少することができる精錬方法に関する。 The present invention relates to a method for refining a copper raw material containing a copper sulfide mineral, and more specifically, in a wet refining method for a copper raw material containing a copper sulfide mineral such as chalcopyrite, with high leaching rate while suppressing sulfur oxidation. The present invention relates to a refining method in which copper is leached and recovered by monovalent copper electrolysis, and accompanying valuable metals are also recovered to reduce waste such as leaching residue as much as possible.
現在、世界の銅の大部分が、銅精鉱を原料とした乾式溶錬法によって製造されている。前記銅精鉱は、黄銅鉱（ＣｕＦｅＳ２）、輝銅鉱（Ｃｕ２Ｓ）、斑銅鉱（Ｃｕ５ＦｅＳ４）など硫化銅鉱物を含有する鉱石を、浮遊選鉱法などの物理分離手段によって硫化鉱物を濃集したものである。前記銅精鉱の組成は、主に鉱石の産地に依存するが、上記した硫化銅鉱物と、黄鉄鉱、磁硫鉄鉱等の硫化鉄鉱物のほか、脈石である珪酸鉱物などの酸化鉱物からなる。また、主に、硫化鉱物中に亜鉛、鉛のほか、ヒ素、アンチモン、ビスマス等のＶ族元素鉱物、セレン、テルル等のＶＩ族元素鉱物、及び貴金属を含有している。 Currently, most of the world's copper is produced by the dry smelting method using copper concentrate as a raw material. The copper concentrate is a sulfide mineral obtained by subjecting ores containing copper sulfide minerals such as chalcopyrite (CuFeS 2 ), chalcocite (Cu 2 S), and chalcopyrite (Cu 5 FeS 4 ) by physical separation means such as a flotation method. Is concentrated. The composition of the copper concentrate mainly depends on the ore production area, but is composed of the above-described copper sulfide mineral, iron sulfide minerals such as pyrite and pyrrhotite, and oxide minerals such as silicate minerals as gangue. In addition to zinc and lead, sulfide minerals mainly contain group V element minerals such as arsenic, antimony and bismuth, group VI element minerals such as selenium and tellurium, and noble metals.
前記乾式溶錬法による銅製錬は、溶錬炉、転炉、精製炉等を用いる一連の乾式製錬で銅精鉱を処理して得られた粗銅を電解精製する方法であり、大量の鉱石を効率よく処理するのに適した方法であるが、その反面、小型設備では反応効率が悪いので、大型設備のために膨大な設備投資が必要であること、また生成する大量のＳＯ２ガスの回収が不可欠であること等の課題がある。 Copper smelting by the above-mentioned dry smelting method is a method of electrolytically purifying crude copper obtained by treating copper concentrate by a series of dry smelting using a smelting furnace, converter, refining furnace, etc. However, on the other hand, since the reaction efficiency is poor in small equipment, a huge equipment investment is required for large equipment, and a large amount of SO 2 gas is generated. There are issues such as the need for collection.
このような状況下、近年、湿式法による精錬方法が研究されている。従来、湿式法による銅精錬としては、銅酸化鉱物を含有する銅鉱石を用いて、積み上げた鉱石に硫酸を散布して銅を浸出し、該浸出生成液の銅濃度を上げるために溶媒抽出法で処理した後、電解採取する方法が工業的に広く用いられている。しかし、銅鉱石の大部分を占める硫化鉱に前記方法を適用した場合、含有鉱物として最も賦存量の多い黄銅鉱では、硫酸による浸出速度が遅く、かつ銅浸出率が低い結果となるという問題があった。したがって、黄銅鉱を含む銅原料の湿式法による精錬方法では、乾式溶錬に匹敵する生産性を得ることは困難であった。 Under such circumstances, in recent years, a refining method using a wet method has been studied. Conventionally, as a copper refining by a wet method, using a copper ore containing a copper oxide mineral, sulfuric acid is sprayed on the piled ore to leach copper, and a solvent extraction method is used to increase the copper concentration of the leaching product liquid. After the treatment with, the method of electrolytic collection is widely used industrially. However, when the above method is applied to sulfide ore occupying the majority of copper ore, chalcopyrite with the largest abundance as a contained mineral has a problem that the leaching rate due to sulfuric acid is slow and the copper leaching rate is low. there were. Therefore, it has been difficult to obtain productivity comparable to dry smelting by a refining method of copper raw material containing chalcopyrite by a wet method.
この解決策として、黄銅鉱の浸出を促進することができる条件で浸出を行う方法が提案されている。代表的な方法として、例えば、銅鉱石又は銅精鉱をハロゲン化物を含む硫酸溶液中で加圧酸化した後浸出し、得られた浸出生成液を溶媒抽出し、第２銅イオンを含む逆抽出液から銅を電解採取する方法（例えば、特許文献１参照）、また、銅精鉱を臭化塩素イオンのようなハロゲン化錯体を形成する浸出液で浸出し、それに続く低酸化還元電位領域での浸出を経て得られた第１銅イオンを含む浸出生成液から銅を電解採取する方法（例えば、特許文献２参照）がある。 As a solution to this problem, a method has been proposed in which leaching is performed under conditions that can promote the leaching of chalcopyrite. As a typical method, for example, copper ore or copper concentrate is subjected to pressure oxidation in a sulfuric acid solution containing a halide and then leached, and the obtained leaching product solution is subjected to solvent extraction and back extraction containing cupric ions. A method of electrolytically collecting copper from the liquid (see, for example, Patent Document 1), and copper concentrate is leached with a leachate that forms a halogenated complex such as a chlorine bromide ion, followed by a low redox potential region. There is a method for electrolytically collecting copper from a leaching product liquid containing first copper ions obtained through leaching (see, for example, Patent Document 2).
上記のような湿式法による精錬方法では、乾式溶錬法に比べて反応温度が低いので、設備が比較的簡便であり投資が圧縮できること、また短周期で運転と停止の繰り返しができるので、生産調整が容易であること等の利点がある。しかし、上記の湿式法に関しても、効率的な精錬方法として、未だ、以下の解決すべき課題がある。 In the refining method by the wet method as described above, the reaction temperature is lower than that of the dry smelting method. Therefore, the equipment is relatively simple and the investment can be compressed. There are advantages such as easy adjustment. However, the above-described wet method still has the following problems to be solved as an efficient refining method.
（１）黄銅鉱での高銅浸出率と硫黄の酸化抑制
湿式法では、硫化鉱物に含まれる硫黄を、硫酸に比べて保管性に優れた元素状硫黄として回収することが望ましい。しかし、難抽出性の黄銅鉱を原料として、銅の高浸出率を得るために強い酸化力を有する浸出法で処理すると、浸出において硫黄が酸化され、浸出液中に硫酸イオンとして溶出し、元素状硫黄として回収することができない。したがって、黄銅鉱を用いて、銅の高浸出率と硫黄の酸化抑制を実現することができる浸出方法が望まれている。
(1) High copper leaching rate in chalcopyrite and oxidation inhibition of sulfur In the wet method, it is desirable to recover sulfur contained in sulfide minerals as elemental sulfur that is superior in storage property compared to sulfuric acid. However, if it is difficult to extract chalcopyrite as a raw material and it is treated by a leaching method with strong oxidizing power to obtain a high leaching rate of copper, sulfur is oxidized in the leaching and eluted as sulfate ions in the leachate. It cannot be recovered as sulfur. Therefore, a leaching method capable of realizing a high copper leaching rate and sulfur oxidation suppression using chalcopyrite is desired.
（２）銅の効率的な還元
湿式法では、第１銅を含む塩化物水溶液から電解採取を行うのが望ましい。すなわち、硫酸溶液では、銅の形態は２価に限られるが、塩化物溶液の場合には、１価と２価の形態がある。一価銅電解、すなわち１価の銅イオンを電解採取する方法では、２価の銅イオンを電解採取する場合の半分の電力量で電解が行えるので経済的である。しかしながら、塩素ガスを用いる浸出工程で得られる浸出生成液では、その酸化還元電位が高いので、銅イオンは２価の形態で存在することになる。従来、第２銅イオンを第１銅イオンに還元する方法において、還元剤として、ＳＯ２ガスを液に吹き込んだり、電解採取で得た銅粉の一部を繰り返すことが知られている。しかし、ＳＯ２の吹き込みでは、浸出液中の硫酸イオン濃度が上昇することになって、液処理のコストが増加してしまう。また、銅粉を使用するとその分製品の直接収率が低下してコスト高につながる。したがって、一価銅電解のため、銅イオンの効率的な還元方法の実現が望まれている。
(2) Efficient reduction of copper In the wet method, it is desirable to perform electrowinning from a chloride aqueous solution containing cuprous. That is, in the sulfuric acid solution, the form of copper is limited to divalent, but in the case of a chloride solution, there are monovalent and divalent forms. Monovalent copper electrolysis, that is, a method of electrolytically collecting monovalent copper ions, is economical because electrolysis can be performed with half the amount of electric power when electrolytically collecting divalent copper ions. However, the leaching product obtained in the leaching process using chlorine gas has a high oxidation-reduction potential, so that copper ions exist in a divalent form. Conventionally, in a method of reducing cupric ions to cuprous ions, it is known that SO 2 gas is blown into the liquid as a reducing agent or a part of copper powder obtained by electrolytic collection is repeated. However, when SO 2 is blown, the concentration of sulfate ions in the leachate increases, and the cost of the liquid treatment increases. Moreover, if copper powder is used, the direct yield of a product will fall correspondingly and it will lead to high cost. Therefore, realization of an efficient reduction method of copper ions is desired for monovalent copper electrolysis.
（３）廃棄残渣量の減少と鉄の効率的回収
湿式法による浸出残渣は、一般に、乾式溶錬によるスラグに比較して化学的に不安定である場合が多く、含まれる不純物が周辺環境に溶出する懸念が指摘されている。この対策として、廃棄する残渣の量を減らすために、浸出において残渣の主成分である鉄の浸出率を上げ、かつ利用可能な形態で鉄を回収することが望ましい。
(3) Reduction of waste residue and efficient recovery of iron In general, leaching residues by wet methods are often more chemically unstable than slag by dry smelting, and impurities contained in the surrounding environment Concerns have been pointed out. As a countermeasure, in order to reduce the amount of residue to be discarded, it is desirable to increase the leaching rate of iron, which is the main component of the residue in leaching, and to recover iron in a usable form.
以上のような状況から、上記の課題を解決して、硫化銅鉱物を含む銅原料を効率的に処理することができる湿式精錬法が望まれている。
なお、本明細書で用いる平均粒子径（Ｄ５０）の測定は、マイクロトラック粒子径分布測定装置（日機装（株）製、型式９３２０ＨＲＡ（Ｘ−１００））を用いて行った。
Under the circumstances as described above, there is a demand for a hydrometallurgical method capable of solving the above-described problems and efficiently treating a copper raw material containing a copper sulfide mineral.
In addition, the measurement of the average particle diameter (D50) used by this specification was performed using the microtrack particle diameter distribution measuring apparatus (The Nikkiso Co., Ltd. make, model 9320HRA (X-100)).
本発明の目的は、上記の従来技術の問題点に鑑み、黄銅鉱を始めとする硫化銅鉱物を含む銅原料の湿式精錬法において、硫黄の酸化を抑制しながら高浸出率で銅を浸出して一価銅電解で回収し、また随伴する有価金属も回収して、浸出残渣などの廃棄物を可能な限り減少することができる精錬方法を提供することにある。 In view of the above-mentioned problems of the prior art, the object of the present invention is to leach copper at a high leaching rate while suppressing sulfur oxidation in a hydrometallurgical process for copper raw materials containing copper sulfide minerals including chalcopyrite. Another object of the present invention is to provide a refining method that can recover waste metal such as leaching residue as much as possible by recovering by monovalent copper electrolysis and also recovering valuable metal.
本発明者らは、上記目的を達成するために、硫化銅鉱物を含む銅原料の精錬方法について、鋭意研究を重ねた結果、硫化銅鉱物を含む銅原料を酸性塩化物水溶液中で塩素で浸出する塩素浸出工程、浸出生成液中の銅イオンを還元する銅イオン還元処理工程、還元生成液中の銅を溶媒で抽出し、次いで抽出液を逆抽出する溶媒抽出工程、銅を電解採取する銅電解採取工程、及び鉄を電解採取する鉄電解採取工程を含む一連のプロセスにより、銅とともに鉄及び随伴する有価金属を効率よく分離回収できることを見出し、本発明を完成するに至った。 In order to achieve the above object, the present inventors have conducted extensive research on a method for refining a copper raw material containing a copper sulfide mineral. As a result, the copper raw material containing a copper sulfide mineral is leached with chlorine in an aqueous acid chloride solution. Chlorine leaching step, copper ion reduction treatment step for reducing copper ions in the leaching product solution, solvent extraction step for extracting the copper in the reduction product solution with a solvent, and then back extracting the extracted solution, copper for electrolytically collecting copper It has been found that iron and accompanying valuable metals can be efficiently separated and recovered together with copper by a series of processes including an electrolytic extraction process and an iron electrolytic extraction process for electrolytically collecting iron, and the present invention has been completed.
すなわち、本発明の第１の発明によれば、硫化銅鉱物を含む銅原料から、湿式精錬法で銅と随伴する有価金属とを回収する方法において、
（１）前記銅原料を酸性塩化物水溶液中で塩素による浸出に付し、該液中に銅を溶出させ、銅イオンを含む浸出生成液と元素状硫黄を含む残渣とを形成する塩素浸出工程、
（２）前記浸出生成液に還元剤を添加し、銅イオンを還元して第１銅イオンを含む還元生成液を得る銅イオン還元処理工程、
（３）前記還元生成液を溶媒抽出に付し、銅を抽出した後、逆抽出して、第１銅イオンを含む逆抽出生成液と第１鉄イオンを含む抽出残液とを得る溶媒抽出工程、
（４）前記逆抽出生成液を電解採取に付し、電着銅と電解尾液とを形成する銅電解採取工程、及び
（５）前記抽出残液を電解採取に付し、電着鉄と塩化鉄水溶液からなる鉄電解尾液とを形成する鉄電解採取工程、を含むことを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。
That is, according to the first invention of the present invention, from a copper raw material containing a copper sulfide mineral, in a method of recovering valuable metals associated with copper by a wet refining method,
(1) Chlorine leaching process in which the copper raw material is subjected to leaching with chlorine in an aqueous acid chloride solution, and copper is eluted in the liquid to form a leaching product liquid containing copper ions and a residue containing elemental sulfur. ,
(2) A copper ion reduction treatment step of adding a reducing agent to the leaching product liquid and reducing the copper ions to obtain a reduced product liquid containing cuprous ions,
(3) Solvent extraction by subjecting the reduction product solution to solvent extraction to extract copper and back-extracting to obtain a back-extraction product solution containing first copper ions and an extraction residual solution containing ferrous ions. Process,
(4) subjecting the back-extraction product solution to electrowinning to form a copper electrowinning step for forming electrodeposited copper and electrolytic tail liquor; and (5) subjecting the extraction residual solution to electrowinning, There is provided a method for refining a copper raw material containing a copper sulfide mineral, comprising an iron electrowinning step of forming an iron electrolytic tail solution comprising an aqueous iron chloride solution.
また、本発明の第２の発明によれば、第１の発明において、塩素浸出工程において、塩素による浸出が塩素ガスの吹きこみによるものであることを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to the second invention of the present invention, in the first invention, in the chlorine leaching process, the leaching with chlorine is caused by blowing in chlorine gas. A refining method is provided.
また、本発明の第３の発明によれば、第１の発明において、塩素浸出工程において、スラリー濃度が１００〜４００ｇ／Ｌ、浸出温度が１００〜１１０℃、酸化還元電位（Ａｇ／ＡｇＣｌ電極規準）が５００〜６００ｍＶであることを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to the third invention of the present invention, in the first invention, in the chlorine leaching step, the slurry concentration is 100 to 400 g / L, the leaching temperature is 100 to 110 ° C., the oxidation-reduction potential (Ag / AgCl electrode standard). ) Is 500 to 600 mV, a method for refining a copper raw material containing a copper sulfide mineral is provided.
また、本発明の第４の発明によれば、第１の発明において、塩素浸出工程において、塩素浸出の終了時点での浸出生成液の塩化物イオン濃度が、２５０〜４００ｇ／Ｌになるように、酸性塩化物水溶液中の塩化物イオン濃度を調整することを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to the fourth invention of the present invention, in the first invention, in the chlorine leaching step, the chloride ion concentration of the leaching product liquid at the end of chlorine leaching is 250 to 400 g / L. There is provided a method for refining a copper raw material containing a copper sulfide mineral, characterized by adjusting a chloride ion concentration in an acidic chloride aqueous solution.
また、本発明の第５の発明によれば、第１の発明において、銅イオン還元処理工程において、還元剤として硫化銅鉱物を用いることを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to a fifth aspect of the present invention, there is provided a method for refining a copper raw material containing a copper sulfide mineral, characterized in that, in the first invention, in the copper ion reduction treatment step, a copper sulfide mineral is used as a reducing agent. Provided.
また、本発明の第６の発明によれば、第５の発明において、前記硫化銅鉱物を浸出生成液中で大気圧下加熱処理に付し、還元生成液と元素状硫黄を含む残渣を形成することを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to a sixth aspect of the present invention, in the fifth aspect, the copper sulfide mineral is subjected to a heat treatment under atmospheric pressure in the leaching product liquid to form a residue containing the reduction product liquid and elemental sulfur. A method for refining a copper raw material containing a copper sulfide mineral is provided.
また、本発明の第７の発明によれば、第６の発明において、前記硫化銅鉱物として平均粒子径（Ｄ５０）が０．５〜６０μｍになるように調製された黄銅鉱を主鉱物とする銅精鉱を用いるとともに、前記浸出生成液の還元温度を９０〜１１０℃の範囲でかつ下記の式１により算出される温度（Ａ）以上に調整することを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。
式１：Ａ（℃）＝６．７９×Ｌｎ（Ｂ）＋８１．５
（式中、Ｂは、銅精鉱の体積頻度累積が５０容量％に相当する平均粒子径（Ｄ５０）を意味し、マイクロトラックを用いて測定された単位μｍの数値である。）
According to a seventh aspect of the present invention, in the sixth aspect, the main mineral is chalcopyrite prepared so that the copper sulfide mineral has an average particle size (D50) of 0.5 to 60 μm. Copper containing copper sulfide mineral, wherein copper concentrate is used, and the reduction temperature of the leaching product liquid is adjusted within the range of 90 to 110 ° C. and higher than the temperature (A) calculated by the following
Formula 1: A (° C.) = 6.79 × Ln (B) +81.5
(In the formula, B means an average particle diameter (D50) in which the volume frequency accumulation of copper concentrate corresponds to 50% by volume, and is a numerical value of unit μm measured using a microtrack.)
また、本発明の第８の発明によれば、第６の発明において、前記硫化銅鉱物として平均粒子径（Ｄ５０）が０．５〜６０μｍになるように調製された黄銅鉱を主鉱物とする銅精鉱を用いるとともに、前記浸出生成液の還元温度を９０〜１１０℃の範囲でかつ下記の式２により算出される温度（Ａ）以上に調整することを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。
式２：Ａ（℃）＝７．０４×Ｌｎ（Ｂ）＋９５．２
（式中、Ｂは、銅精鉱の体積頻度累積が５０容量％に相当する平均粒子径（Ｄ５０）を意味し、マイクロトラックを用いて測定された単位μｍの数値である。）
According to the eighth invention of the present invention, in the sixth invention, the main mineral is chalcopyrite prepared so that the copper sulfide mineral has an average particle diameter (D50) of 0.5 to 60 μm. Copper containing copper sulfide mineral, wherein copper concentrate is used, and the reduction temperature of the leaching product liquid is adjusted in the range of 90 to 110 ° C. and higher than the temperature (A) calculated by the following
Formula 2: A (° C.) = 7.04 × Ln (B) +95.2
(In the formula, B means an average particle diameter (D50) in which the volume frequency accumulation of copper concentrate corresponds to 50% by volume, and is a numerical value of unit μm measured using a microtrack.)
また、本発明の第９の発明によれば、第６の発明において、前記残渣を銅原料として前記塩素浸出工程に送ることを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to a ninth aspect of the present invention, there is provided a method for refining a copper raw material containing a copper sulfide mineral in the sixth aspect, wherein the residue is sent to the chlorine leaching step as a copper raw material. .
また、本発明の第１０の発明によれば、第１の発明において、溶媒抽出工程において、溶媒抽出に用いられる有機溶媒が中性抽出剤を含むことを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to a tenth aspect of the present invention, in the first aspect, in the solvent extraction step, an organic solvent used for solvent extraction contains a neutral extractant, and a copper raw material containing a copper sulfide mineral A refining method is provided.
また、本発明の第１１の発明によれば、第１０の発明において、前記有機溶媒中の中性抽出剤の濃度が、４０容量％以上であることを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to an eleventh aspect of the present invention, in the tenth aspect, the concentration of the neutral extractant in the organic solvent is 40% by volume or more, and the copper raw material containing a copper sulfide mineral A refining method is provided.
また、本発明の第１２の発明によれば、第１の発明において、溶媒抽出工程において、逆抽出で用いられる水溶液は、銅濃度が７０ｇ／Ｌ以下、塩素イオン濃度が５０〜３５０ｇ／Ｌであることを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to the twelfth aspect of the present invention, in the first aspect, the aqueous solution used in the back extraction in the solvent extraction step has a copper concentration of 70 g / L or less and a chlorine ion concentration of 50 to 350 g / L. There is provided a method for refining a copper raw material containing a copper sulfide mineral.
また、本発明の第１３の発明によれば、第１の発明において、溶媒抽出工程において、逆抽出の温度が、４０〜９０℃であることを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to the thirteenth aspect of the present invention, in the first aspect, in the solvent extraction step, the temperature of back extraction is 40 to 90 ° C. A method is provided.
また、本発明の第１４の発明によれば、第１の発明において、銅電解採取工程において、陰極室、陽極室、及び前記両室を分離する隔膜から構成される電解槽を用いて、該陰極室に前記溶媒抽出工程からの第１銅イオンを含む逆抽出生成液を給液して銅を電析させ、かつ該陽極室に前記鉄電解採取工程からの塩化鉄水溶液からなる鉄電解尾液を給液して陽極酸化させるとともに、該陽極室への給液が隔膜を通じて該陰極室へ流入するのを防止することを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to a fourteenth aspect of the present invention, in the first aspect, in the copper electrowinning step, an electrolytic cell comprising a cathode chamber, an anode chamber, and a diaphragm separating the two chambers is used. A negative extraction product solution containing cuprous ions from the solvent extraction step is supplied to the cathode chamber to deposit copper, and an iron electrolytic tail comprising an aqueous iron chloride solution from the iron electrowinning step is supplied to the anode chamber. Provided is a method for refining a copper raw material containing a copper sulfide mineral, wherein the liquid is supplied and anodized to prevent the liquid supplied to the anode chamber from flowing into the cathode chamber through a diaphragm. .
また、本発明の第１５の発明によれば、第１４の発明において、前記隔膜の通水度が、０．０４〜０．１５Ｌ／ｍ２．ｓであることを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to a fifteenth aspect of the present invention, in the fourteenth aspect, the water permeability of the diaphragm is 0.04 to 0.15 L / m 2 . A method for refining a copper raw material containing a copper sulfide mineral is provided.
また、本発明の第１６の発明によれば、第１４の発明において、前記電解槽は、給液と廃液が陰極室と陽極室のそれぞれで個別に行われ、かつ陰極室の液面レベルを陽極室のそれよりも高くする構造であることを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to a sixteenth aspect of the present invention, in the fourteenth aspect, the electrolytic cell is configured such that the supply liquid and the waste liquid are separately performed in each of the cathode chamber and the anode chamber, and the liquid level of the cathode chamber is set. Provided is a method for refining a copper raw material containing a copper sulfide mineral, characterized in that the structure is higher than that of an anode chamber.
また、本発明の第１７の発明によれば、第１４の発明において、前記陰極室の廃液を溶媒抽出の逆抽出液として前記溶媒抽出工程へ戻すとともに、前記陽極室の廃液を浸出液として前記塩素浸出工程へ戻すことを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to a seventeenth aspect of the present invention, in the fourteenth aspect, the cathode chamber waste liquid is returned to the solvent extraction step as a back extraction liquid for solvent extraction, and the anode chamber waste liquid is used as the leachate to produce the chlorine. Provided is a method for refining a copper raw material containing a copper sulfide mineral, characterized by returning to a leaching step.
また、本発明の第１８の発明によれば、第１の発明において、鉄電解採取工程において、濾布で仕切られた陽極室と陰極室から構成される電解槽内で、陽極室に陰極で析出する鉄量の２倍量以上の鉄イオンを供給して浴電圧を低下させて電解を行うことを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to an eighteenth aspect of the present invention, in the first aspect, in the iron electrowinning process, the anode chamber is made of a cathode in an electrolytic cell composed of an anode chamber and a cathode chamber partitioned by a filter cloth. Provided is a method for refining a copper raw material containing a copper sulfide mineral, characterized in that an iron ion having an amount twice or more the amount of precipitated iron is supplied to lower a bath voltage to perform electrolysis.
また、本発明の第１９の発明によれば、第１の発明において、鉄電解採取工程に先立って、溶媒抽出工程で得られる抽出残液を浄液に付し、精製液と沈殿生成物とを形成する浄液工程を含むことを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to the nineteenth aspect of the present invention, in the first aspect, prior to the iron electrowinning step, the extraction residual liquid obtained in the solvent extraction step is applied to the purified solution, and the purified solution and the precipitated product There is provided a method for refining a copper raw material containing a copper sulfide mineral, characterized in that it comprises a liquid purification step for forming a copper sulfide.
また、本発明の第２０の発明によれば、第１９の発明において、前記浄液工程において、浄液が、硫化処理、セメンテーション処理、又は中和処理から選ばれる少なくとも１種であることを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to a twentieth aspect of the present invention, in the nineteenth aspect, in the liquid purification step, the liquid purification is at least one selected from sulfurization treatment, cementation treatment, or neutralization treatment. A method for refining a copper raw material containing the copper sulfide mineral is provided.
また、本発明の第２１の発明によれば、第１の発明において、さらに、塩素浸出工程で得られる元素状硫黄を含む残渣を、不活性雰囲気下で蒸留に付し、硫黄を揮発させ、凝縮された硫黄と随伴する貴金属を含む残滓とに分離する浸出残渣処理工程を含むことを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 According to the twenty-first aspect of the present invention, in the first aspect, the residue containing elemental sulfur obtained in the chlorine leaching step is subjected to distillation under an inert atmosphere to volatilize the sulfur, There is provided a method for refining a copper raw material containing a copper sulfide mineral, characterized by including a leach residue treatment step of separating into condensed sulfur and a residue containing a precious metal accompanying it.
また、本発明の第２２の発明によれば、第１の発明において、さらに、銅電解採取工程で得られる電着銅を陽極として、電解精製に付し、高純度の電気銅と銀含有スライムとに分離する銅電解精製工程を含むことを特徴とする硫化銅鉱物を含む銅原料の精錬方法が提供される。 Further, according to the twenty-second invention of the present invention, in the first invention, the electrodeposited copper obtained in the copper electrowinning step is used as an anode and subjected to electrolytic purification to obtain high-purity electrolytic copper and silver-containing slime. The copper refining method containing the copper sulfide mineral characterized by including the copper electrolytic refining process isolate | separated into these is provided.
本発明の硫化銅鉱物を含む銅原料の精錬方法は、黄銅鉱を始めとする硫化銅鉱物を含む銅原料の湿式精錬法において、硫黄の酸化を抑制しながら高浸出率で銅を浸出して一価銅電解で回収することができ、また随伴する有価金属も回収して、浸出残渣などの廃棄物を可能な限り減少することができる精錬方法であり、その工業的価値は極めて大きい。 The method for refining a copper raw material containing a copper sulfide mineral according to the present invention is a method for refining copper raw material containing a copper sulfide mineral such as chalcopyrite, and leaching copper at a high leaching rate while suppressing oxidation of sulfur. This is a refining method that can be recovered by monovalent copper electrolysis and can also recover the accompanying valuable metals to reduce waste such as leach residue as much as possible, and its industrial value is extremely high.
以下、本発明の硫化銅鉱物を含む銅原料の精錬方法を詳細に説明する。
まず、本発明の硫化銅鉱物を含む銅原料の精錬方法の概要について、図を用いて説明する。図１は、本発明の精錬プロセス工程図の一例を表す。
図１において、硫化銅鉱物を含む銅原料８は、最初に塩素浸出工程１に付され、銅、鉄等を含有する浸出生成液と硫黄含有残渣とに分離される。浸出生成液は、銅イオン還元処理工程２に付され、浸出生成液中の銅イオンは還元され、第１銅イオンを含む還元生成液が得られる。ここで、還元剤として硫化銅鉱物を含む銅原料を用いる場合は、この残渣は塩素浸出工程１へ循環される。還元生成液は、溶媒抽出工程３に付され、溶媒抽出及び逆抽出により第１銅イオンを含有する逆抽出生成液と抽出残液に分離される。逆抽出生成液は、銅電解採取工程４に付され、銅は電着銅９として回収される。
Hereinafter, the refining method of the copper raw material containing the copper sulfide mineral of the present invention will be described in detail.
First, the outline | summary of the refining method of the copper raw material containing the copper sulfide mineral of this invention is demonstrated using figures. FIG. 1 shows an example of a refining process flow chart of the present invention.
In FIG. 1, a copper
また、精錬処理の原料の種類にもよるが、通常硫化銅鉱物を含む銅鉱石は、銅とほぼ同量に近い鉄を含有しており、前記溶媒抽出工程３における抽出残液には、多量の鉄イオンが含まれる。したがって、溶媒抽出工程３における抽出残液は、必要に応じて浄液工程５に付され、鉄イオン含有精製液と鉄以外の有価金属固形物とに分離される。鉄イオン含有精製液は、鉄電解採取工程６に付され、鉄は電着鉄１０として回収される。
Further, although depending on the type of raw material for the refining treatment, the copper ore containing the copper sulfide mineral usually contains iron that is close to the same amount as copper, and the extraction residual liquid in the
また、塩素浸出工程１で分離された硫黄含有残渣は浸出残渣処理工程７に付され、元素状硫黄が回収される。さらに、銅電解採取工程４で分離された電解尾液は、陰極電解尾液が逆抽出給液として溶媒抽出工程３に、陽極電解尾液が浸出液として塩素浸出工程１に再循環される。また、鉄電解採取工程６で得られる電解尾液は陽極給液として銅電解採取工程４へ送られる。
Further, the sulfur-containing residue separated in the
１．硫化銅鉱物を含む銅原料と随伴する有価金属
本発明の精錬方法における硫化銅鉱物を含む銅原料としては、黄銅鉱（ＣｕＦｅＳ２）、輝銅鉱（Ｃｕ２Ｓ）、斑銅鉱（Ｃｕ５ＦｅＳ４）などの硫化銅鉱物を含む銅鉱石、前記銅鉱石から浮遊選鉱法等によって硫化銅鉱物を濃集した銅精鉱、硫化銅鉱物を含み、酸化銅鉱物、ヒ化銅鉱物、アンチモン化銅鉱物など各種含銅鉱物を含む鉱石及びその銅精鉱、並びに銅精鉱などから乾式溶錬法で得られる銅マットおよび高品位銅マットが含まれ、さらには、これらと同時処理される硫化物状、酸化物状、金属状の各種含銅原料がある場合も含まれる。
1. Copper raw materials containing copper sulfide minerals and valuable metals accompanying the copper raw materials containing copper sulfide minerals in the refining method of the present invention include chalcopyrite (CuFeS 2 ), chalcocite (Cu 2 S), and chalcopyrite (Cu 5 FeS 4). ) And other copper ores containing copper sulfide minerals, copper concentrates concentrated from the copper ore by the flotation method, copper sulfide minerals, copper oxide minerals, copper arsenide minerals, antimony copper minerals Ores containing various copper-containing minerals such as copper concentrates, copper mats obtained by dry smelting from copper concentrates and high-grade copper mats, and sulfides that are treated simultaneously with these In addition, there are cases where there are various copper-containing raw materials in the form of oxides and metals.
また、随伴する有価金属としては、鉄、ニッケル、コバルト、マンガン、硫黄、亜鉛、カドミウム、錫、鉛の他、ヒ素、アンチモン、ビスマス等のＶ族元素、セレン、テルル等のＶＩ族元素、及び貴金属等が挙げられる。 The accompanying valuable metals include iron, nickel, cobalt, manganese, sulfur, zinc, cadmium, tin, lead, group V elements such as arsenic, antimony and bismuth, group VI elements such as selenium and tellurium, and A noble metal etc. are mentioned.
２．塩素浸出工程
本発明の精錬方法における塩素浸出工程は、上記硫化銅鉱物を含む銅原料を塩化銅、塩化鉄などを含む酸性塩化物水溶液中に懸濁させ、主に硫化銅鉱物を塩素で浸出して銅、鉄等を溶出させて、銅イオンと鉄イオンを含む浸出生成液と元素状硫黄を含む残渣とを形成する工程である。また、上記銅原料を次工程の銅イオン還元処理工程で還元剤として用いる場合には、該工程で得られる残渣を塩素浸出工程の原料として用いることが好ましい。例えば、上記銅原料の全量を一旦銅イオン還元処理工程で処理した後に、塩素浸出工程で用いることができる。
2. Chlorine leaching step The chlorine leaching step in the refining method of the present invention involves suspending the copper raw material containing the copper sulfide mineral in an acidic chloride aqueous solution containing copper chloride, iron chloride, etc., and leaching mainly the copper sulfide mineral with chlorine. Then, copper, iron and the like are eluted to form a leaching product liquid containing copper ions and iron ions and a residue containing elemental sulfur. Moreover, when using the said copper raw material as a reducing agent in the copper ion reduction process process of the following process, it is preferable to use the residue obtained at this process as a raw material of a chlorine leaching process. For example, the entire amount of the copper raw material can be used in the chlorine leaching step after once being treated in the copper ion reduction treatment step.
上記工程において浸出に用いる塩素は、特に限定されるものではなく、酸性塩化物水溶液中の塩化銅、塩化鉄などを酸化することができる塩素化合物が用いられるが、この中で、特に反応効率から、酸化力の強い塩素ガスの吹きこみを行うのが好ましい。 Chlorine used for leaching in the above process is not particularly limited, and a chlorine compound that can oxidize copper chloride, iron chloride, etc. in an aqueous solution of acidic chloride is used. It is preferable to blow in chlorine gas having strong oxidizing power.
上記工程において、輝銅鉱、斑銅鉱及び高品位銅マットに比べて、塩素による浸出反応の速度が遅い黄銅鉱が主たる浸出対象である場合には、銅の高浸出率とともに硫黄の酸化抑制が重要な課題である。この課題に対応するためには、浸出時の酸化力を適正に制御することが望ましい。なお、酸化力は、浸出液の酸化還元電位、浸出温度、スラリー濃度、浸出液の塩素濃度等の条件を最適化することによって制御することができる。 In the above process, when chalcopyrite, which has a slower rate of leaching reaction with chlorine compared to chalcocite, porphyry and high-grade copper matte, is the main leaching target, it is important to suppress sulfur oxidation along with the high leaching rate of copper. It is a difficult task. In order to cope with this problem, it is desirable to appropriately control the oxidizing power during leaching. The oxidizing power can be controlled by optimizing conditions such as the redox potential of the leachate, the leach temperature, the slurry concentration, and the chlorine concentration of the leachate.
上記工程における塩素浸出液の酸化還元電位（ＯＲＰと呼称することがある。Ａｇ／ＡｇＣｌ電極規準）は、特に限定されるものではなく、５００〜６００ｍＶが好ましく、５００〜５２０ｍＶがより好ましい。すなわち、ＯＲＰが５００ｍＶ未満では、浸出の酸化力が弱いため、銅の浸出率が低い。一方、６００ｍＶを超えて浸出すると、硫黄の酸化率が著しく増加する。さらに、ＯＲＰが５００〜５２０ｍＶでは硫黄の酸化はほとんど生じない。また、銅原料が黄銅鉱主体の原料である場合には、ＯＲＰは５００〜５２０ｍＶが特に好ましい。 The oxidation-reduction potential of the chlorine leachate in the above step (sometimes referred to as ORP. Ag / AgCl electrode standard) is not particularly limited, and is preferably 500 to 600 mV, more preferably 500 to 520 mV. That is, when the ORP is less than 500 mV, the leaching oxidizing power is weak, so the copper leaching rate is low. On the other hand, when leaching exceeds 600 mV, the oxidation rate of sulfur increases remarkably. Furthermore, when the ORP is 500 to 520 mV, sulfur oxidation hardly occurs. Moreover, when the copper raw material is a raw material mainly composed of chalcopyrite, the ORP is particularly preferably 500 to 520 mV.
上記工程における浸出温度は、特に限定されるものではなく、１００〜１１０℃が好ましく、より好ましくは１０５〜１１０℃である。すなわち、浸出温度が１００℃未満では、銅及び鉄の浸出率が低く、その反面硫黄は酸化される。一方、１１０℃を超えると加圧設備が必要となる。 The leaching temperature in the above step is not particularly limited, and is preferably 100 to 110 ° C, more preferably 105 to 110 ° C. That is, when the leaching temperature is less than 100 ° C., the leaching rate of copper and iron is low, and on the other hand, sulfur is oxidized. On the other hand, if it exceeds 110 degreeC, a pressurization installation will be needed.
上記工程における浸出初期のスラリー濃度（スラリー中の銅原料の濃度）は、特に限定されるものではなく、１００〜４００ｇ／Ｌが好ましく、より好ましくは２５０〜４００ｇ／Ｌである。すなわち、浸出初期のスラリー濃度が１００ｇ／Ｌ未満では、銅及び鉄の浸出率が低く、その反面硫黄は酸化される。一方、４００ｇ／Ｌを超えると、設備及び操作上のむずかしさがある。 The slurry concentration (concentration of the copper raw material in the slurry) at the beginning of leaching in the above step is not particularly limited, and is preferably 100 to 400 g / L, more preferably 250 to 400 g / L. That is, when the slurry concentration at the initial stage of leaching is less than 100 g / L, the leaching rate of copper and iron is low, and on the other hand, sulfur is oxidized. On the other hand, when it exceeds 400 g / L, there is difficulty in equipment and operation.
さらに、上記工程における浸出終了時点での浸出終液に含まれる塩素イオン濃度は、特に限定されるものではなく、２００〜４００ｇ／Ｌが好ましく、より好ましくは２５０〜４００ｇ／Ｌである。すなわち、浸出終了時点での浸出終液に含まれる塩素イオン濃度が、２００ｇ／Ｌ未満では、銅及び鉄の浸出率が低く、その反面硫黄は酸化される。一方、４００ｇ／Ｌを超えてもそれ以上の反応結果への効果がない。浸出終了時点での浸出終液に含まれる塩素イオン濃度を２００〜４００ｇ／Ｌになるように維持することによって、黄銅鉱中の銅をほぼ完全に浸出させることができる。なお、浸出終液に含まれる塩素イオン濃度を２００〜４００ｇ／Ｌに維持するためには、酸性塩化物水溶液への塩素吹き込み量を制御する方法が好ましい。 Furthermore, the chlorine ion concentration contained in the leaching final solution at the end of leaching in the above step is not particularly limited, and is preferably 200 to 400 g / L, more preferably 250 to 400 g / L. That is, when the chlorine ion concentration contained in the final leaching solution at the end of leaching is less than 200 g / L, the leaching rate of copper and iron is low, and on the other hand, sulfur is oxidized. On the other hand, even if it exceeds 400 g / L, there is no effect on the reaction result. By maintaining the chlorine ion concentration in the leaching final solution at the end of leaching to be 200 to 400 g / L, copper in chalcopyrite can be almost completely leached. In addition, in order to maintain the chlorine ion concentration contained in the leaching final solution at 200 to 400 g / L, a method of controlling the amount of chlorine blown into the acidic chloride aqueous solution is preferable.
本発明の精錬方法において、塩素浸出工程を上記の条件で行うことにより、黄銅鉱を主体とする銅原料を用いて、黄銅鉱の硫黄の酸化率を５％以下に抑制しつつ、銅の９５％以上と鉄の９０％以上を浸出することができる。これによって、湿式銅精錬プロセスの課題の一つである黄銅鉱の高銅浸出率と硫黄の酸化抑制が達成される。
また、塩素浸出工程において形成される第２銅イオン及び第２鉄イオンを含む浸出生成液と元素状硫黄を含む残渣は、通常の手段によって固液分離される。
In the refining method of the present invention, the chlorine leaching step is performed under the above-described conditions, thereby using a copper raw material mainly composed of chalcopyrite and suppressing the oxidation rate of chalcopyrite sulfur to 5% or less, while maintaining 95% of copper. % And over 90% of iron can be leached. As a result, the high copper leaching rate of chalcopyrite and the suppression of oxidation of sulfur, which are one of the problems of the wet copper refining process, are achieved.
In addition, the leaching product liquid containing cupric ions and ferric ions formed in the chlorine leaching step and the residue containing elemental sulfur are solid-liquid separated by ordinary means.
３．銅イオン還元処理工程
本発明の精錬方法における銅イオン還元処理工程は、上記塩素浸出工程で得られる銅イオン、鉄イオン等を含有する浸出生成液に還元剤を添加して銅イオンの還元処理を行い、浸出生成液に含有される第２銅イオンを第１銅イオンに還元し、同時に第２鉄イオンも第１鉄イオンに還元する工程である。これによって得られる第１銅イオンが高比率で存在する還元生成液から、次の溶媒抽出工程において、銅イオンのみを選択的に有機溶媒に抽出させることができる。
3. Copper ion reduction treatment step The copper ion reduction treatment step in the refining method of the present invention is a reduction treatment of copper ions by adding a reducing agent to the leaching product liquid containing copper ions, iron ions, etc. obtained in the chlorine leaching step. This is a step of reducing the cupric ions contained in the leaching product liquid to cuprous ions and simultaneously reducing the ferric ions to ferrous ions. In the next solvent extraction step, only the copper ions can be selectively extracted into the organic solvent from the reduction product liquid in which the first copper ions are present in a high ratio.
上記工程において、還元生成液のＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）は、銅と鉄を含む塩化物水溶液中の第２銅イオンを第１銅イオンへ還元することができる電位に調整されるが、０〜４００ｍＶが好ましく、０〜３８０ｍＶがより好ましい。すなわち、ＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）が４００ｍＶを超えると、銅イオンの一部は２価となり、さらにこの第２銅イオンが酸化剤として働いて鉄イオンも一部３価の状態となるので、第１銅イオンが高比率で存在する還元生成液が得られない。一方、ＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）が０ｍＶ未満であると、場合によって鉄イオン又は銅イオンが金属状態まで還元されて沈殿することがある。 In the above process, the ORP (Ag / AgCl electrode standard) of the reduction product solution is adjusted to a potential that can reduce the cupric ion in the aqueous chloride solution containing copper and iron to the cuprous ion. 0 to 400 mV is preferable, and 0 to 380 mV is more preferable. That is, when the ORP (Ag / AgCl electrode standard) exceeds 400 mV, some of the copper ions become divalent, and further, the cupric ions act as an oxidizing agent, and the iron ions also become trivalent. A reduction product solution containing cuprous ions in a high ratio cannot be obtained. On the other hand, if the ORP (Ag / AgCl electrode standard) is less than 0 mV, iron ions or copper ions may be reduced to a metal state and precipitate in some cases.
ところで、鉄イオンなどが共存する浸出生成液の第１銅と第２銅の形態を正確に分析することは困難であるが、図２より、銅の形態とＯＲＰの関係を推定した。
図２は、試薬を混合して、９０℃で塩化物溶液中の第１鉄イオン濃度を変えた場合の銅の形態（Ｃｕ（１）／Ｃｕ（２）：第１銅と第２銅の濃度比）とＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）の関係を示す。ここで、銅及び鉄イオンの大部分が第１銅イオンと第１鉄イオンで存在するには、ＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）が４００ｍＶ以下、完全に第１銅イオンと第１鉄イオンとするには、３８０ｍＶ以下であることが分る。したがって、Ｃｕ濃度が５０ｇ／Ｌ程度で５０〜１００ｇ／Ｌの濃度でＦｅが共存する浸出生成液の場合には、概ね３８０〜４００ｍＶが還元工程での到達目標のＯＲＰとなる。
By the way, although it is difficult to accurately analyze the form of the cuprous and cupric copper in the leaching product liquid in which iron ions coexist, the relationship between the form of copper and the ORP was estimated from FIG.
Fig. 2 shows the form of copper (Cu (1) / Cu (2): cuprous and cupric copper) when the ferrous ion concentration in the chloride solution is changed at 90 ° C by mixing the reagents. The relationship between the concentration ratio) and ORP (Ag / AgCl electrode standard) is shown. Here, in order for most of copper and iron ions to exist as cuprous ions and ferrous ions, the ORP (Ag / AgCl electrode standard) is 400 mV or less, and the cuprous ions and ferrous ions are completely It turns out that it is 380 mV or less. Therefore, in the case of a leaching product liquid in which Fe coexists at a Cu concentration of about 50 g / L and a concentration of 50 to 100 g / L, approximately 380 to 400 mV is the ORP that is the target reached in the reduction step.
上記工程で用いる還元剤としては、特に限定されるものではなく、金属銅、銅よりも卑な金属、硫化銅又は硫化銅鉱物から選ばれる少なくとも１種が使用できるが、例えば、特に銅の浸出も同時に行える硫化銅又は硫化銅鉱物を用いるのが好ましい。すなわち、上記塩素浸出工程に先立って、硫化銅鉱物を含む銅原料を塩素浸出工程からの浸出生成液と予め接触させることによって、銅イオンを１価に、鉄イオンを２価に還元し、同時に原料中の銅の一部を予め浸出できる。このとき、形成される元素状硫黄を含む残渣は、銅原料として上記塩素浸出工程に送られ浸出に付されることが好ましい。 The reducing agent used in the above step is not particularly limited, and at least one selected from metallic copper, a base metal rather than copper, copper sulfide, or copper sulfide mineral can be used. It is preferable to use copper sulfide or copper sulfide mineral that can be used simultaneously. That is, prior to the chlorine leaching step, a copper raw material containing a copper sulfide mineral is brought into contact with a leaching product solution from the chlorine leaching step in advance to reduce copper ions to monovalent and iron ions to bivalent, A part of copper in the raw material can be leached in advance. At this time, it is preferable that the residue containing elemental sulfur to be formed is sent to the chlorine leaching step as a copper raw material and subjected to leaching.
ここで、硫化銅鉱物が黄銅鉱の場合に、第２銅イオンと第２鉄イオンを、第１銅イオンと第１鉄イオンに還元し、かつ元素状の硫黄を生成する還元反応は、以下の化学反応式１及び２で表わせる。
Here, when the copper sulfide mineral is chalcopyrite, the reduction reaction for reducing cupric ions and ferric ions to cuprous ions and ferrous ions and generating elemental sulfur is as follows. These can be represented by
化学反応式１：Ｃｕ２＋＋１／３ＣｕＦｅＳ２→４／３Ｃｕ＋＋１／３Ｆｅ２＋＋２／３Ｓ、
化学反応式２：Ｆｅ３＋＋１／３ＣｕＦｅＳ２→１／３Ｃｕ＋＋４／３Ｆｅ２＋＋２／３Ｓ
Chemical reaction formula 1: Cu 2+ + 1 / 3CuFeS 2 → 4 / 3Cu + + 1 / 3Fe 2+ + 2 / 3S,
Chemical reaction formula 2: Fe 3+ + 1 / 3CuFeS 2 → 1 / 3Cu + + 4 / 3Fe 2+ + 2 / 3S
化学反応式１及び２の反応は、第２銅イオンと第２鉄イオンによる黄銅鉱の浸出反応である。したがって、第２銅イオンと第２鉄イオンの還元反応に伴ない黄銅鉱の浸出が進む。
The reaction of
上記工程において、還元剤として硫化銅鉱物を用いる場合、還元処理条件は、特に限定されるものではなく、大気圧下又は加圧下で行われる。この中で、特に加圧設備が不要な大気圧下での還元処理が好ましい。すなわち、硫化銅鉱物を浸出生成液中で大気圧下加熱処理に付し、還元生成液と元素状硫黄を含む残渣を形成する方法が好ましい。 In the above step, when copper sulfide mineral is used as the reducing agent, the reduction treatment conditions are not particularly limited, and are performed under atmospheric pressure or under pressure. Of these, reduction treatment under atmospheric pressure that does not require pressurization equipment is particularly preferable. That is, a method of subjecting the copper sulfide mineral to a heat treatment under atmospheric pressure in the leaching product solution to form a residue containing the reduction product solution and elemental sulfur is preferable.
前記大気圧下での還元処理方法としては、特に限定されるものではなく、例えば、以下の二つの方法で、第１銅イオンが高比率で存在する還元状態が得られる。
すなわち、一つは、前記硫化銅鉱物として平均粒子径（Ｄ５０）が０．５〜６０μｍになるように調製された黄銅鉱を主鉱物とする銅精鉱を用いるとともに、前記浸出生成液の還元温度を９０〜１１０℃の範囲でかつ下記の式１により算出される温度（Ａ）以上に調整することを含む方法である。これによって、還元生成液のＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）を４００ｍＶ以下とすることができる。
式１：Ａ（℃）＝６．７９×Ｌｎ（Ｂ）＋８１．５
（式中、Ｂは、銅精鉱の体積頻度累積が５０容量％に相当する平均粒子径（Ｄ５０）を意味し、マイクロトラックを用いて測定された単位μｍの数値である。）
The reduction treatment method under atmospheric pressure is not particularly limited. For example, a reduced state in which cuprous ions are present in a high ratio can be obtained by the following two methods.
That is, one uses a copper concentrate whose main mineral is chalcopyrite prepared so that an average particle diameter (D50) is 0.5 to 60 μm as the copper sulfide mineral, and reduction of the leaching product liquid. This is a method including adjusting the temperature within the range of 90 to 110 ° C. and higher than the temperature (A) calculated by the following
Formula 1: A (° C.) = 6.79 × Ln (B) +81.5
(In the formula, B means an average particle diameter (D50) in which the volume frequency accumulation of copper concentrate corresponds to 50% by volume, and is a numerical value of unit μm measured using a microtrack.)
また、二つめは、前記硫化銅鉱物として平均粒子径（Ｄ５０）が０．５〜６０μｍになるように調製された黄銅鉱を主鉱物とする銅精鉱を用いるとともに、浸出生成液の還元温度を９０〜１１０℃の範囲でかつ下記の式２により算出される温度（Ａ）以上に調整することを含む方法である。これによって、還元生成液のＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）を３８０ｍＶ以下とすることができる。
式２：Ａ（℃）＝７．０４×Ｌｎ（Ｂ）＋９５．２
（式中、Ｂは、銅精鉱の体積頻度累積が５０容量％に相当する平均粒子径（Ｄ５０）を意味し、マイクロトラックを用いて測定された単位μｍの数値である。）
The second is a copper concentrate using chalcopyrite as the main mineral and having a mean particle size (D50) of 0.5 to 60 μm as the copper sulfide mineral, and the reduction temperature of the leaching solution. Is adjusted in the range of 90 to 110 ° C. and higher than the temperature (A) calculated by the following
Formula 2: A (° C.) = 7.04 × Ln (B) +95.2
(In the formula, B means an average particle diameter (D50) in which the volume frequency accumulation of copper concentrate corresponds to 50% by volume, and is a numerical value of unit μm measured using a microtrack.)
前記二つの方法において、浸出生成液中の第２銅イオン及び第２鉄イオンと、黄銅鉱との反応性が、還元生成液の酸化還元電位の低下、すなわち還元反応の進行度合にとって重要である。このため、硫化銅鉱物として所定の平均粒子径（Ｄ５０）に粒子径が調整された銅精鉱を用いることと、前記銅精鉱のＤ５０から式１又は２に従って求められる温度から還元温度を適正に選択することの二つの要件が必須である。
In the two methods, the reactivity of the cupric and ferric ions in the leaching solution and chalcopyrite is important for the reduction of the redox potential of the reduction solution, that is, the degree of progress of the reduction reaction. . For this reason, the copper concentrate whose particle diameter is adjusted to a predetermined average particle diameter (D50) is used as the copper sulfide mineral, and the reduction temperature is set appropriately from the temperature determined according to the
これによって、銅精鉱の粒子径を所定値に調整して用いれば、浸出生成液を大気圧下で所定の還元温度に加熱するとき、ＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）を第１銅イオンが第２銅イオンに対して優位に存在する４００ｍＶ以下、好ましくは３８０ｍＶ以下に低下させることができる。すなわち、大気圧下の処理で第２銅イオンを第１銅イオンに効率的に還元することができることを意味する。 Thus, if the particle diameter of the copper concentrate is adjusted to a predetermined value and used, when the leaching product liquid is heated to a predetermined reduction temperature under atmospheric pressure, the ORP (Ag / AgCl electrode standard) is changed by the first copper ion. It can be lowered to 400 mV or less, preferably 380 mV or less, which exists predominantly with respect to cupric ions. That is, it means that cupric ions can be efficiently reduced to cuprous ions by treatment under atmospheric pressure.
上記還元方法で用いる銅精鉱の粒子径は、Ｄ５０が０．５〜６０μｍであり、好ましくは１〜１０μｍ、さらに好ましくは３．５〜８．２μｍである。すなわち、Ｄ５０が０．５μｍ未満では、還元生成液と残渣との分離工程での漏れ又は沈降性の悪化等が起る。一方６０μｍを超えると、銅精鉱の反応性が低下するため、到達目標のＯＲＰを得るためには浸出生成液の温度を上げてより沸点に近い温度で長時間保持することになるので、生産性及びエネルギー的に非効率である。 As for the particle diameter of the copper concentrate used by the said reduction method, D50 is 0.5-60 micrometers, Preferably it is 1-10 micrometers, More preferably, it is 3.5-8.2 micrometers. That is, when D50 is less than 0.5 μm, leakage in the separation step of the reduction product solution and the residue, deterioration of sedimentation, or the like occurs. On the other hand, if it exceeds 60 μm, the reactivity of copper concentrate will decrease, so in order to obtain the target ORP, the temperature of the leaching product solution will be raised and kept at a temperature close to the boiling point for a long time. It is inefficient and energetically inefficient.
上記還元方法で用いる銅精鉱の粒子径の調整には、特に限定されるものではなく、市販の各種の粉砕機、例えば、通常ビーズミルと呼ばれる湿式媒体撹拌粉砕機（例えば浅田鉄工製ナノグレンミルＮＭ−Ｇ５Ｍ、アメックス製ＮＶＭ−２）及び遊星ボールミル（例えばセイシン企業製プラネタリーミルＳＦＫ−０４）等が用いられる。 Adjustment of the particle diameter of the copper concentrate used in the reduction method is not particularly limited, and various commercially available grinders, for example, a wet medium agitating grinder usually called a bead mill (for example, Nanogren Mill NM- manufactured by Asada Tekko) G5M, Amex NVM-2), planetary ball mill (for example, planetary mill SFK-04 manufactured by Seishin Corporation) and the like are used.
上記還元方法で用いる還元温度の範囲は、９０〜１１０℃である。すなわち、上記塩素浸出工程では、例えば、浸出液の塩化物イオン濃度は２００〜４００ｇ／Ｌが好ましいことから、このときの浸出生成液の沸点は１１０℃近傍まで上昇する。ここで、還元温度が１１０℃を超えると大気圧下での処理ができない。一方９０℃未満では、上記銅精鉱の粒子径を細かくしても反応が遅く、目標のＯＲＰを得るためには長時間の処理を要するので効率が低い。 The range of the reduction temperature used in the reduction method is 90 to 110 ° C. That is, in the chlorine leaching step, for example, the chloride ion concentration of the leaching solution is preferably 200 to 400 g / L, so that the boiling point of the leaching product solution at this time rises to around 110 ° C. Here, when the reduction temperature exceeds 110 ° C., the treatment under atmospheric pressure cannot be performed. On the other hand, when the temperature is less than 90 ° C., the reaction is slow even if the particle size of the copper concentrate is reduced, and the efficiency is low because a long time treatment is required to obtain the target ORP.
上記還元方法で用いる銅精鉱の初期スラリー濃度は、特に限定されるものではないが、高濃度であるほど還元反応が進み易いが、プロセスの物量バランス、塩化銅及び塩化鉄の溶解度等から制約がある。この中で、操業上の制約が少ない５０〜２５０ｇ／Ｌが望ましく、１００ｇ／Ｌがより望ましい。すなわち、前記初期スラリー濃度が１００ｇ／Ｌであれば、実用上の問題はない。 The initial slurry concentration of the copper concentrate used in the above reduction method is not particularly limited, but the higher the concentration, the easier the reduction reaction proceeds, but there are limitations due to the balance of process quantities, the solubility of copper chloride and iron chloride, etc. There is. In this, 50-250 g / L with few restrictions on operation is desirable, and 100 g / L is more desirable. That is, if the initial slurry concentration is 100 g / L, there is no practical problem.
上記還元方法において、銅精鉱の平均粒子径（Ｄ５０）と適正な還元温度の関係は、銅精鉱のＤ５０（Ｂ）から下記の式１又は２を満足する温度（Ａ）を求め、前記浸出生成液の還元温度を９０〜１１０℃の範囲でかつ温度（Ａ）以上に調整する。
式１：Ａ（℃）＝６．７９×Ｌｎ（Ｂ）＋８１．５
式２：Ａ（℃）＝７．０４×Ｌｎ（Ｂ）＋９５．２
（式中、Ｂは、銅精鉱の体積頻度累積が５０容量％に相当する平均粒子径（Ｄ５０）を意味し、マイクロトラックを用いて測定された単位μｍの数値である。）
これによって、還元生成液のＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）を、式１に従えば４００ｍＶ以下に、また式２に従えば３８０ｍＶ以下にすることができる。すなわち、第１銅イオンが高比率で存在する還元生成液を得ることができる。
In the above reduction method, the relationship between the average particle diameter (D50) of copper concentrate and the appropriate reduction temperature is determined from the D50 (B) of copper concentrate to the temperature (A) satisfying the
Formula 1: A (° C.) = 6.79 × Ln (B) +81.5
Formula 2: A (° C.) = 7.04 × Ln (B) +95.2
(In the formula, B means an average particle diameter (D50) in which the volume frequency accumulation of copper concentrate corresponds to 50% by volume, and is a numerical value of unit μm measured using a microtrack.)
Thereby, the ORP (Ag / AgCl electrode standard) of the reduction product liquid can be reduced to 400 mV or less according to
ここで、前記の式１又は２は、銅精鉱の初期スラリー濃度１００ｇ／Ｌで、それぞれ４００ｍＶ以下又は３８０ｍＶ以下の目標ＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）を達成するための銅精鉱のＤ５０と還元温度の関係を示すもので、以下によって導かれた。まず、図３に、還元温度と粒子径を変えて還元処理を行ったときの還元温度、粒子径、ＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）の関係を示す。図中に、９０、１０４、１０７℃の温度での近似線を示す。これより得られた目標ＯＲＰと温度及びＤ５０との関係を表１、図４及び図５に示す。図４と図５の近似線の回帰式を、各々式１と式２とした。
Here, the
以上より明らかなように、安価な黄銅鉱を主鉱物とする銅精鉱を用いて第２銅イオンを還元して、第１銅イオンが高比率で存在する還元生成液を得ることができる。 As is clear from the above, it is possible to obtain a reduction product liquid in which the cuprous ions are present in a high ratio by reducing cupric ions using copper concentrate containing cheap chalcopyrite as the main mineral.
以上、本発明の精錬方法における銅イオン還元処理工程によって、塩化物水溶液中の第２銅イオンと第２鉄イオンを効率良く還元し、かつ硫化銅鉱物中の硫黄の酸化を抑制して反応を進めて、第１銅イオン及び第１鉄イオンを含む還元生成液と元素状硫黄を含む残渣とを形成することができる。これによって、湿式銅精錬プロセスの課題の一つである銅の効率的な還元が達成される。 As described above, the copper ion reduction treatment step in the refining method of the present invention efficiently reduces the cupric ions and ferric ions in the aqueous chloride solution, and suppresses the oxidation of sulfur in the copper sulfide mineral to react. Proceeding, a reduction product liquid containing cuprous ions and ferrous ions and a residue containing elemental sulfur can be formed. This achieves efficient reduction of copper, which is one of the problems of the wet copper refining process.
４．溶媒抽出工程
本発明の精錬方法における溶媒抽出工程は、上記銅イオン還元処理工程で得られる第１銅イオンを含む還元生成液と有機抽出剤を含む有機溶媒とを接触混合させて第１銅イオンのみを選択的に有機溶媒に抽出する工程と、第１銅イオンを抽出した有機溶媒と水溶液とを接触混合させて、第１銅イオンを水溶液に逆抽出する工程とによって、銅イオン含有水溶液と鉄イオン及び有価金属イオン含有抽出残液を得る工程である。なお、前記還元生成液は、ＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）が０〜４００ｍＶに調整されているものである。
4). Solvent extraction step The solvent extraction step in the refining method of the present invention comprises a reduction product solution containing the first copper ion obtained in the copper ion reduction treatment step and an organic solvent containing an organic extractant in contact with each other to obtain a first copper ion. A step of selectively extracting the first copper ions into an organic solvent, and a step of contacting and mixing the organic solvent from which the first copper ions have been extracted and an aqueous solution, and back extracting the first copper ions into the aqueous solution, This is a step of obtaining an extraction residue containing iron ions and valuable metal ions. The reduction product liquid is one in which ORP (Ag / AgCl electrode standard) is adjusted to 0 to 400 mV.
上記工程に用いる有機抽出剤は、特に限定されるものではなく、第１銅イオンを抽出し、鉄及び随伴する有価金属と分離できる有機抽出剤であれば、いずれも用いることができるが、この中で、特にトリブチルフォスフェイトなどの中性抽出剤が好ましい。すなわち、トリブチルフォスフェイト抽出剤を用い、還元生成液の酸化還元電位を最適に維持して溶媒抽出することによって、第１銅イオンを有機溶媒相に選択的に抽出し、第１鉄イオン及び銀イオンなどを抽出残液に残すことができる。 The organic extractant used in the above step is not particularly limited, and any organic extractant that can extract cuprous ions and separate them from iron and accompanying valuable metals can be used. Among these, a neutral extractant such as tributyl phosphate is particularly preferable. That is, by using a tributyl phosphate extractant and performing solvent extraction while optimally maintaining the redox potential of the reduction product solution, the cuprous ions are selectively extracted into the organic solvent phase, and the ferrous ions and silver are extracted. Ions etc. can be left in the extraction residual liquid.
ここで、有機溶媒相中のトリブチルフォスフェイトなどの中性抽出剤の濃度は、特に限定されるものではないが、好ましくは４０〜１００容量％、さらに好ましくは５０〜１００容量％である。すなわち、４０容量％未満では、工業的に期待する銅抽出率を得ることができない。通常、トリブチルフォスフェイトは流動性を保つためにケロシン等の希釈剤で希釈して用いられるが、銅イオンの抽出率の向上には、トリブチルフォスフェイトの希釈は極力行わない方が望ましい。すなわち、前記銅イオンの抽出率は、還元生成液中の塩化物イオンの濃度と、トリブチルフォスフェイトの濃度とに依存するからである。 Here, the concentration of the neutral extractant such as tributyl phosphate in the organic solvent phase is not particularly limited, but is preferably 40 to 100% by volume, more preferably 50 to 100% by volume. That is, if it is less than 40 volume%, the copper extraction rate expected industrially cannot be obtained. In general, tributyl phosphate is diluted with a diluent such as kerosene in order to maintain fluidity. However, it is desirable not to dilute tributyl phosphate as much as possible in order to improve the extraction rate of copper ions. That is, the extraction rate of the copper ions depends on the concentration of chloride ions and the concentration of tributyl phosphate in the reduction product solution.
上記工程の逆抽出に用いる水溶液の銅濃度は、特に限定されるものではないが、７０ｇ／Ｌ以下が好ましい。すなわち、前記銅濃度が７０ｇ／Ｌを超えると逆に有機溶媒相に銅が移動する現象が起こる。 Although the copper concentration of the aqueous solution used for the back extraction of the said process is not specifically limited, 70 g / L or less is preferable. That is, when the copper concentration exceeds 70 g / L, a phenomenon occurs in which copper moves to the organic solvent phase.
上記工程の逆抽出に用いる水溶液の塩素イオン濃度は、特に限定されるものではないが、５０〜３５０ｇ／Ｌが好ましい。すなわち、前記塩素イオン濃度が５０ｇ／Ｌ未満では、逆抽出される第１銅イオンは水への溶解度が小さいので、逆抽出された銅イオンを溶液の状態に保つことができない。通常、逆抽出される銅濃度にあわせて、逆抽出に用いる溶液の塩素イオン濃度を高くして、逆抽出された銅イオンを溶液の状態に保つが、実用的には、塩素イオン濃度の上限は３５０ｇ／Ｌであるので、この値が塩素イオン濃度の上限となる。 Although the chlorine ion concentration of the aqueous solution used for the back extraction of the said process is not specifically limited, 50-350 g / L is preferable. That is, when the chlorine ion concentration is less than 50 g / L, the back-extracted first copper ions have low solubility in water, and therefore the back-extracted copper ions cannot be kept in a solution state. Usually, in accordance with the copper concentration to be back-extracted, the concentration of chlorine ions in the solution used for back-extraction is increased and the back-extracted copper ions are kept in the solution state. Is 350 g / L, so this value is the upper limit of the chlorine ion concentration.
上記工程の逆抽出の温度は、特に限定されるものではないが、４０〜９０℃が好ましく、より好ましくは５０〜９０℃である。すなわち、４０℃未満では、トリブチルフォスフェイト中の銅イオンは水相側へ排出されにくく、逆抽出率が小さい。一方、９０℃を超えると、放熱量が多くなり、温度を保つことが困難になるうえ、溶媒の蒸散量も多くなって有機溶媒相及び水相を安定な状態を保つことができない。 Although the temperature of the back extraction of the said process is not specifically limited, 40-90 degreeC is preferable, More preferably, it is 50-90 degreeC. That is, when the temperature is lower than 40 ° C., copper ions in tributyl phosphate are not easily discharged to the aqueous phase side, and the back extraction rate is small. On the other hand, if it exceeds 90 ° C., the amount of heat release increases and it becomes difficult to maintain the temperature, and the amount of transpiration of the solvent increases and the organic solvent phase and the aqueous phase cannot be kept stable.
以上、本発明の精錬方法における溶媒抽出工程によって、銅イオン還元処理工程で得られた還元生成液から、第１銅イオン含有水溶液と鉄イオン及び有価金属イオン含有抽出残液が効率よく得られる。 As described above, by the solvent extraction step in the refining method of the present invention, the first copper ion-containing aqueous solution, the iron ion and valuable metal ion-containing extraction residual liquid can be efficiently obtained from the reduction product obtained in the copper ion reduction treatment step.
５．銅電解採取工程
本発明の精錬方法における銅電解採取工程は、上記溶媒抽出工程で得られる第１銅イオンを含む逆抽出生成液から銅を電解採取し、陰極上に析出された電着銅と電解尾液とを形成する工程である。
5). Copper electrowinning step The copper electrowinning step in the refining method of the present invention is obtained by electrolytically collecting copper from the back extraction product solution containing the first copper ions obtained in the solvent extraction step, and depositing copper on the cathode. This is a step of forming an electrolytic tail solution.
上記工程において、銅の電解採取方法は、特に限定されるものではなく、塩化物からの電解採取法により金属を回収する種々の方法を用いることができるが、この中で、特に、陰極室、陽極室、及び前記両室を分離する隔膜から構成される電解槽を用いて、該陰極室に溶媒抽出工程からの逆抽出生成液（塩化第１銅水溶液）を給液して銅を電析させ、かつ該陽極室に鉄電解採取工程からの鉄電解尾液（塩化鉄水溶液）を給液して陽極酸化させるとともに、該陽極室への給液が隔膜を通じて該陰極室へ流入するのを防止することを含む隔膜電解による方法を用いるのが、好ましい。さらに、上記方法で前記陰極室の廃液を溶媒抽出の逆抽出液として溶媒抽出工程へ戻すとともに、前記陽極室の廃液を浸出液として塩素浸出工程へ戻すのが、さらに好ましい。 In the above step, the method of electrolytic collection of copper is not particularly limited, and various methods of recovering metal by electrolytic collection from chloride can be used. Using an electrolytic cell composed of an anode chamber and a diaphragm separating the two chambers, the cathode chamber is fed with a back-extraction product solution (cuprous chloride aqueous solution) from the solvent extraction step to deposit copper. And anodizing by supplying an iron electrolytic tail solution (iron chloride aqueous solution) from the iron electrowinning process to the anode chamber, and supplying the anode chamber through the diaphragm to the cathode chamber. It is preferable to use a method by diaphragm electrolysis that includes preventing. Furthermore, it is more preferable that the waste liquid in the cathode chamber is returned to the solvent extraction step as a back extraction liquid for solvent extraction by the above method, and the waste liquid in the anode chamber is returned to the chlorine leaching step as a leachate.
上記工程の銅の電解採取方法を、図を用いて説明する。図６は、本発明の銅電解採取工程で用いる電解槽の構造の一例を表わす図である。
図６の電解槽は、隔膜１３によって陰極室１１と陽極室１２に別けられ、それぞれの電極が設置される。陰極１４は、特に限定されるものでなく、金属銅、チタン、ステンレスが使用できる。また、陽極１５は、特に限定されるものでなく、食塩電解等の塩化物水溶液から塩素ガス発生用に用いる不溶性電極、例えば、商品名ＤＳＥ（ペルメレック電極（株）製）が使用できる。
The method for electrolytically collecting copper in the above process will be described with reference to the drawings. FIG. 6 is a diagram showing an example of the structure of an electrolytic cell used in the copper electrowinning process of the present invention.
The electrolytic cell in FIG. 6 is divided into a cathode chamber 11 and an
また、前記電解槽の特徴は、陽極室１２への給液１９が隔膜１３を通じて陰極室１１へ流入するのを防止する構造であることである。このため、陰極室液２０と陽極室液２１とが分離されるが、隔膜１３を通してイオン及び電気が通過する必要があるので、陰極室液２０と陽極室液２１を厳格に分割するものではない。すなわち、陰極室１１に酸化された陽極室液２１が自由に流入しない構造であればよく、イオン及び水の通過を完全に停止するものである必要はない。
The electrolytic cell is characterized in that it has a structure that prevents the liquid 19 supplied to the
これを電解槽の構造上で実現するため、隔膜は、特に限定されるものではなく、例えば濾布又は固体電解質膜が用いられるが、この中でも、特に目が細かく、通水度が低くなるように織られた濾布を用いる方法が好ましい。すなわち、固体電解質膜は、濾布と比べてコストが高く、また不純物に弱いからである。 In order to realize this on the structure of the electrolytic cell, the diaphragm is not particularly limited. For example, a filter cloth or a solid electrolyte membrane is used, but among these, the eyes are particularly fine and the water permeability is low. A method using a filter cloth woven into the fabric is preferred. That is, the solid electrolyte membrane is more expensive than the filter cloth and is vulnerable to impurities.
前記隔膜の通水度は、特に限定されるものではないが、０．０４〜０．１５Ｌ／ｍ２．ｓが好ましい。すなわち、０．０４Ｌ／ｍ２．ｓ未満では、液の移動が少ないため槽電圧が上昇し、また濾布のコストも上昇する。一方０．１５Ｌ／ｍ２．ｓを超えれば、液の移動の増加により銅の収率が低下する。 The water permeability of the diaphragm is not particularly limited, but is 0.04 to 0.15 L / m 2 . s is preferred. That is, 0.04 L / m 2 . If it is less than s, since the movement of the liquid is small, the cell voltage increases and the cost of the filter cloth also increases. On the other hand, 0.15 L / m 2 . If s is exceeded, the yield of copper will fall by the increase in movement of a liquid.
さらに、陽極室１２に陰極室液２０が少量流入して陽極室液２１に混入しても操業上影響がほとんどなく問題がない場合には、陽極室液２１が陰極室１１側に流入しないようにするために陰極室液２０の水頭を陽極室液２１より高く保つことが好ましい。例えば、陰極室１１のオーバーフローレベルを陽極室１２のそれよりも若干高めとし、液面差を付けてその圧力で陽極室１２の塩素ガスや塩素ガスを含んだ液が陰極室１１に入り込まないようにする。
Further, if there is no problem in operation even if a small amount of the
ここで、電解採取を順調に進行させるため、陰極での還元反応と陽極での酸化反応（陽極酸化）からなる電解反応を平衡させることが必要である。
そのため、陰極室１１には、上記溶媒抽出工程からの逆抽出生成液（塩化第１銅水溶液）を陰極給液１８として供給して、陰極１４上で第１銅イオンを金属銅に還元し電析させる。反応後の陰極室液２０は、陰極室１１から直接排出され、回収されるような構造を持つ装置とする。
Here, it is necessary to equilibrate the electrolytic reaction composed of the reduction reaction at the cathode and the oxidation reaction (anodic oxidation) at the anode in order to proceed the electrolytic collection smoothly.
Therefore, the cathode chamber 11 is supplied with the back-extraction product solution (cuprous chloride aqueous solution) from the solvent extraction step as the
一方、陽極室１２には、鉄電解採取工程からの鉄電解尾液（塩化鉄水溶液）を陽極給液１９として供給して、陽極１５上で塩化鉄水溶液を陽極酸化させる。反応後の陽極室液２１は、陽極室１２から直接排出され、回収されるような構造を持つ装置とする。ここで、鉄電解採取工程からの塩化鉄水溶液は、陽極で酸化されて電子を放出することができるイオンを含んだ溶液であるので電解反応が成り立つ。例えば、鉄電解採取工程で、浴電圧を下げるために陽極反応として塩化第１鉄を塩化第２鉄に酸化する反応を用いる場合、鉄電解陽極廃液は塩素イオンを充分に含む塩化第２鉄溶液であり、銅電解採取の陽極給液１９として好ましい。この塩化第２鉄水溶液を銅電解採取工程で陽極酸化すると塩素ガスが生成される。
On the other hand, an iron electrolytic tail solution (iron chloride aqueous solution) from the iron electrowinning step is supplied to the
また、隔膜１３は陰極室液２０と陽極室液２１を厳格に分割するものではないので、陰極室液２０の一部が陽極室液２１に混入して排出されるが、大部分の陰極室液２０はそのまま陰極廃液２２として排出されることになる。これにより、陰極室液２０中の電着しなかった第１銅イオンは、酸化性の陽極室液２１によって酸化されることなく回収することができる。このため、溶媒抽出工程の逆抽出液として反応面の問題がなく好適である。
Further, since the diaphragm 13 does not strictly divide the
また、陽極廃液２３は、塩素浸出工程に循環されるが、銅電解採取工程での液量の増減が殆どないため、液バランス上の問題は起こらない。また、酸化剤である塩化第２鉄が更に酸化力のある塩素ガスに形を変えることになり、浸出工程での効率向上に好ましい。
Further, the
陰極廃液２２及び陽極廃液２３の排出方法としては、各室の液面を一定に保つことが出来れば良く、それぞれの室の液面にあわせたオーバーフロー口１６、１７を設け、自動的に排出される機構とするのが簡便な方法である。
As a method for discharging the
以上、本発明の精錬方法における銅電解採取工程では、第１銅イオンの電解採取が行われるので高電流効率が達成される。また、そこで得られる陰極室液は還元性であるので溶媒抽出の逆抽出液として反応的に安定かつ好適な性状であり、かつ陽極室液は強い酸化性であるので塩化浸出工程の浸出液として好適な性状である。さらに、これらの液を循環することによって、循環される工程を含むプロセス全体の液量バランスをとることができる。 As described above, in the copper electrowinning step in the refining method of the present invention, high current efficiency is achieved because the electrowinning of the first copper ion is performed. Moreover, the cathodic chamber liquid obtained therein is reducible, so that it is reactively stable and suitable as a back extraction liquid for solvent extraction, and the anodic chamber liquid is highly oxidative, so it is suitable as a leaching liquid for the chlorination leaching process. It is a characteristic. Furthermore, by circulating these liquids, it is possible to balance the liquid amount of the entire process including the steps to be circulated.
６．鉄電解採取工程
本発明の精錬方法における鉄電解採取工程は、上記溶媒抽出工程で得られる抽出残液から鉄を電解採取して、陰極に析出された電着鉄と上記銅電解採取工程に好適な陽極給液を形成する工程である。電着鉄は、純鉄あるいは屑鉄として利用することができるので、鉄電解採取工程は残渣処理場所と資源の有効利用の点から好ましい。
6). Iron electrowinning step The iron electrowinning step in the refining method of the present invention is suitable for the electrodeposited iron deposited on the cathode and the copper electrowinning step by electrowinning iron from the extraction residual liquid obtained in the solvent extraction step. This is a step of forming a positive anode liquid. Since electrodeposited iron can be used as pure iron or scrap iron, the iron electrowinning process is preferred from the standpoint of residue treatment and effective utilization of resources.
上記工程における鉄の電解採取方法は、特に限定されるものではないが、例えば、隔膜電解法を用いて、前記抽出残液を電解槽の陰極給液とし、陽極室から陽極廃液を得るのが好ましい。前記陽極廃液は塩化鉄水溶液であるので、新たに銅を浸出するための浸出液として塩素浸出工程へ循環することができるが、銅電解採取工程の陽極給液として循環するのが好ましい。すなわち、銅電解採取工程を経由して陽極給液として作用させた後に塩素浸出工程へ循環されるのが、湿式銅精錬プロセス全体の効率において特に好ましい。 The method for electrolytically collecting iron in the above step is not particularly limited. For example, by using a diaphragm electrolysis method, the extraction residual liquid is used as a cathode supply liquid in an electrolytic cell, and an anode waste liquid is obtained from an anode chamber. preferable. Since the anode waste solution is an aqueous iron chloride solution, it can be circulated to the chlorine leaching step as a leaching solution for newly leaching copper, but it is preferably circulated as an anode feed solution for the copper electrowinning step. That is, it is particularly preferable in terms of the efficiency of the entire wet copper refining process that it is made to act as an anode feed solution via the copper electrowinning step and then circulated to the chlorine leaching step.
上記の隔膜電解法としては、特に限定されるものではないが、塩素ガスの発生抑制と浴電圧の低下のために、特定の給液方法及び電解条件で行う隔膜電解法を用いるのが好ましい。すなわち、陽極では、陰極で析出する鉄イオンの２倍量の２価の鉄イオンが３価に酸化されるので、電解採取する鉄量を陰極に給液した２価の鉄イオンの３分の１以下に制御する。ここで、電解反応が進行するためには、陽極で鉄イオンが２価から３価に酸化されるだけで十分であるので、塩素ガスの発生が抑制される。したがって、隔膜電解法で、陽極室に陰極で析出する鉄量の２倍量以上の鉄イオンを供給することで、塩素ガスの発生を抑制して浴電圧を低下させることができる。 The diaphragm electrolysis method is not particularly limited, but it is preferable to use a diaphragm electrolysis method performed under a specific liquid supply method and electrolysis conditions in order to suppress the generation of chlorine gas and lower the bath voltage. That is, at the anode, bivalent iron ions that are twice as much as the iron ions deposited at the cathode are oxidized to trivalent, so the amount of iron to be electrolyzed is 3/3 of the divalent iron ions fed to the cathode. Control to 1 or less. Here, in order for the electrolytic reaction to proceed, it is sufficient that the iron ions are oxidized from divalent to trivalent at the anode, so that generation of chlorine gas is suppressed. Therefore, by supplying the iron ion more than twice the amount of iron deposited at the cathode to the anode chamber by the diaphragm electrolysis method, generation of chlorine gas can be suppressed and the bath voltage can be lowered.
なお、溶媒抽出工程又はその後の処理中に液中の２価の鉄イオンが３価に酸化された場合には、鉄電解採取工程での電流効率を低下させて電力コストを上昇させるので、鉄電解採取工程の給液口に鉄粉あるいは鉄板等を設置することで液中の３価の鉄イオンを２価に還元するのが好ましい。 If divalent iron ions in the liquid are oxidized to trivalent during the solvent extraction step or subsequent processing, the current efficiency in the iron electrowinning step is reduced and the power cost is increased. It is preferable to reduce trivalent iron ions in the liquid to divalent by installing iron powder or an iron plate or the like at the liquid supply port in the electrolytic collection step.
以上、鉄電解採取工程における鉄の電解採取方法として特定の給液方法及び電解条件で行う上記隔膜電解法を用いれば、電着鉄と上記銅電解採取工程に好適な陽極給液を形成するとともに、塩素ガスの発生を抑制して浴電圧を低下させることができる。これによって、湿式銅精錬プロセスの課題の一つである廃棄残渣量の減少と鉄の効率的回収が達成される。 As described above, if the diaphragm electrolysis method performed under a specific liquid supply method and electrolysis conditions is used as an iron electrowinning method in the iron electrowinning step, an electrode feed suitable for the electrodeposited iron and the copper electrowinning step is formed. The generation of chlorine gas can be suppressed and the bath voltage can be lowered. As a result, reduction of the amount of waste residue and efficient recovery of iron, which are one of the problems of the wet copper refining process, are achieved.
７．浄液工程
本発明の精錬方法は、必要に応じて、鉄電解採取工程に先立って、溶媒抽出工程で得られる鉄を含む抽出残液を処理する浄液工程を行うことができる。浄液工程は、溶媒抽出工程で得られた鉄を含む抽出残液から、随伴する有価金属を沈殿生成させ分離し、随伴する有価金属を含む固形物と鉄を含む精製液とを分離する工程である。
一般に、鉄の電解採取においては、鉄は他の不純物と電位的にも共析し易いことから資源として利用できる鉄を得るためには不純物の除去が課題となる。したがって、上記溶媒抽出工程で得られた鉄を含む抽出残液から、随伴する有価金属を回収し、また資源として利用できる鉄を得るために、浄液工程を設けることができる。
7). Liquid Purification Process The refining method of the present invention can perform a liquid purification process for treating the extraction residual liquid containing iron obtained in the solvent extraction process prior to the iron electrowinning process, if necessary. The liquid purification step is a step of separating and separating the accompanying valuable metal from the extraction residual liquid containing iron obtained in the solvent extraction step, and separating the solid containing the accompanying valuable metal and the purified liquid containing iron. It is.
In general, in the electrowinning of iron, removal of impurities becomes a problem in order to obtain iron that can be used as a resource because iron is likely to eutect with other impurities in terms of potential. Therefore, a liquid purification step can be provided in order to recover the accompanying valuable metals from the extraction residue containing iron obtained in the solvent extraction step and to obtain iron that can be used as a resource.
浄液工程には、従来公知の方法の適用が可能であるが、この中でも硫化処理、セメンテーション処理又は中和処理から選ばれる少なくとも１種の処理方法が好ましく、随伴する有価金属の種類、含有量によって適宜選択することができる。なお、随伴する有価金属の種類及び含有量は、原料中の含有状態に左右され、また溶媒抽出の条件によっても変化するので限定されない。
なお、浄液によって形成された随伴する有価金属を含む固形物と鉄を含む精製液は、通常の手段によって固液分離される。
Conventionally known methods can be applied to the liquid purification step. Among these, at least one treatment method selected from sulfidation treatment, cementation treatment or neutralization treatment is preferred, and the type and content of the accompanying valuable metals It can be appropriately selected depending on the amount. The type and content of the accompanying valuable metal are not limited because they depend on the content in the raw material and vary depending on the solvent extraction conditions.
In addition, the solid containing the valuable metal and the refined liquid containing iron formed by the liquid purification are separated into solid and liquid by ordinary means.
８．浸出残渣処理工程
本発明の精錬方法は、必要に応じて、塩素浸出工程で得られる元素状硫黄を含む残渣を処理する浸出残渣処理工程を行うことができる。浸出残渣処理工程は、塩素浸出工程で得られる元素状硫黄を含む残渣を不活性雰囲気下で加熱して蒸留処理し、硫黄を揮発させ、凝縮された硫黄と随伴する貴金属類を含む残滓とに分離する工程である。
8). Leaching residue treatment step The refining method of the present invention can perform a leaching residue treatment step of treating the residue containing elemental sulfur obtained in the chlorine leaching step, as necessary. In the leaching residue treatment step, the residue containing elemental sulfur obtained in the chlorine leaching step is heated and distilled in an inert atmosphere to volatilize the sulfur, to form a residue containing condensed sulfur and associated precious metals. It is a process of separating.
前記蒸留処理の温度は、特に限定されるものではないが、２５０〜３５０℃が好ましく、３３０〜３５０℃がより好ましい。すなわち、２５０℃未満では、硫黄の揮発率が低い。
ここで、前記蒸留処理において蒸留装置の冷却部から凝縮された硫黄を得て、必要に応じて、さらに精製して硫黄製品とすることができる。一方、貴金属類は、残滓中に濃縮され、既存の製錬、精製工程の貴金属回収法で処理することができる。なお、得られた残滓は、元素状硫黄が完全に除かれているので、含有される貴金属が効率良く回収される。
Although the temperature of the said distillation process is not specifically limited, 250-350 degreeC is preferable and 330-350 degreeC is more preferable. That is, at a temperature lower than 250 ° C., the volatility of sulfur is low.
Here, sulfur condensed from the cooling unit of the distillation apparatus in the distillation treatment can be obtained and further refined as necessary to obtain a sulfur product. On the other hand, the noble metals are concentrated in the residue and can be processed by the noble metal recovery method in the existing smelting and refining processes. In addition, since the elemental sulfur is completely removed from the obtained residue, the precious metal contained is efficiently recovered.
９．銅電解精製工程
本発明の精錬方法は、必要に応じて、銅電解採取工程で得られる電着銅を処理する銅電解精製工程を行うことができる。銅電解精製工程は、銅電解採取工程で得られる電着銅を陽極として、電解精製に付し、高純度の電気銅と銀含有スライムとに分離する工程である。銅電解採取工程で得られる電着銅の不純物濃度が高い場合、必要により、銅電解精製工程を行う。例えば、前記電着銅を溶融して陽極を鋳造し、この陽極を通常の電解精製法で処理して、不純物及び貴金属をスライム又は電解液へ分配させて、陰極上に高純度な電気銅を得ることができる。また、前記電着銅を、既存の乾式溶錬法の後半工程の転炉ないし精製炉に投入して陽極を鋳造し、この陽極を既存の銅電解精製工程で処理することもできる。
9. Copper electrolytic refining process The refining method of this invention can perform the copper electrolytic refining process of processing the electrodeposited copper obtained by a copper electrowinning process as needed. The copper electrolytic purification step is a step of subjecting the electrodeposited copper obtained in the copper electrowinning step to electrolytic purification and separating it into high-purity electrolytic copper and silver-containing slime. When the impurity concentration of the electrodeposited copper obtained in the copper electrowinning process is high, a copper electrolytic purification process is performed as necessary. For example, the electrodeposited copper is melted to cast an anode, and the anode is treated by a usual electrolytic purification method to distribute impurities and noble metals to slime or electrolyte, and high purity electrolytic copper is formed on the cathode. Can be obtained. Moreover, the electrodeposited copper can be put into a converter or refining furnace in the latter half of the existing dry smelting method to cast an anode, and this anode can be processed in the existing copper electrolytic refining process.
以上、本発明に係わるプロセスの工程について説明したが、その構成は、銅原料の組成（硫化銅の種類構成、鉄濃度、貴金属含有量、他の有価物含有量など）、回収製品の品質、工場立地などによって、選ぶことが出来る。例えば、鉄含有量が少なくその大部分が浸出残渣として系外排出されるような場合、特に輝銅鉱（Ｃｕ２Ｓ）や高品位銅マットが原料であるときには、溶媒抽出工程、浄液工程、及び鉄回収工程を省略できる。 As mentioned above, although the process of the process according to the present invention has been described, the composition thereof is the composition of the copper raw material (type composition of copper sulfide, iron concentration, noble metal content, content of other valuable materials, etc.), the quality of the recovered product, You can choose according to the factory location. For example, when the iron content is low and most of it is discharged out of the system as a leaching residue, particularly when chalcopyrite (Cu 2 S) or high-grade copper mat is a raw material, a solvent extraction step, a liquid purification step, And the iron recovery process can be omitted.
以下、本発明の実施例によって本発明を詳細に説明するが、本発明はこれらの実施例によって何ら限定されるものではない。なお、実施例で用いた分析方法及び平均粒子径（Ｄ５０）の測定方法は以下の通りである。
（１）金属の分析：液体試料はそのまま、固形試料は酸溶解して、ＩＣＰ発光分析法で行った。
（２）塩素イオン濃度の分析：硝酸銀を用いた電位差滴定で塩酸濃度を分析し、ＩＣＰ発光分析法で求めた銅、鉄及びナトリウムイオンに付随する塩素イオン濃度を算出し、これを合算した。
（３）鉱物種組成及び硫黄形態の分析：顕微鏡観察により鉱物種を同定し、化学分析値から推定した。
（４）平均粒子径（Ｄ５０）の測定：マイクロトラック粒子径分布測定装置（日機装（株）製、型式９３２０ＨＲＡ（Ｘ−１００））を用いて行った。
EXAMPLES Hereinafter, although the present invention will be described in detail by examples of the present invention, the present invention is not limited to these examples. In addition, the measuring method used in the Example and the measuring method of an average particle diameter (D50) are as follows.
(1) Analysis of metal: The liquid sample was used as it was, the solid sample was dissolved in an acid, and the ICP emission analysis was performed.
(2) Chlorine ion concentration analysis: The hydrochloric acid concentration was analyzed by potentiometric titration using silver nitrate, and the chloride ion concentrations associated with the copper, iron and sodium ions determined by ICP emission analysis were calculated and added.
(3) Analysis of mineral species composition and sulfur form: Mineral species were identified by microscopic observation and estimated from chemical analysis values.
(4) Measurement of average particle size (D50): Measurement was performed using a microtrack particle size distribution measuring apparatus (manufactured by Nikkiso Co., Ltd., model 9320HRA (X-100)).
（実施例１）
（１）塩素浸出工程
銅原料を酸性塩化物水溶液中で塩素による浸出に付し、この液中に銅を溶出させ、銅イオンを含む浸出生成液と元素状硫黄を含む残渣とを得て、評価した。
銅原料として、表２に示す化学組成で、かつ表３に示す鉱物種組成である銅精鉱を使用した。
(Example 1)
(1) Chlorine leaching step The copper raw material is subjected to leaching with chlorine in an aqueous acid chloride solution, and copper is eluted in this liquid to obtain a leaching product liquid containing copper ions and a residue containing elemental sulfur. evaluated.
As a copper raw material, a copper concentrate having a chemical composition shown in Table 2 and a mineral species composition shown in Table 3 was used.
銅精鉱３０〜１２０ｇの所定量を秤量し、銅濃度６０ｇ／Ｌ及び塩素イオン濃度２００ｇ／Ｌの酸性塩化物水溶液３００ｍＬとともに、容量５００ｍＬのチタン製の反応容器に装入した。このスラリー濃度を１００〜４００ｇ／Ｌとした。
オイルバスを使用して、前記容器内を１０５〜１１０℃の温度に維持し、ＯＲＰをＡｇ／ＡｇＣｌ電極を参照電極として４５０〜７５０ｍＶの間の所定の一定値に維持するように、塩素ガスを吹き込んで塩素浸出処理を行った。処理開始から、１、３、６時間経過後にサンプリングし、浸出残渣中の存在物量と浸出液中の銅イオン濃度、鉄イオン濃度及び硫黄濃度を分析し、銅精鉱からの銅と鉄の浸出率及び硫黄酸化率を算出した。
A predetermined amount of 30 to 120 g of copper concentrate was weighed and charged into a titanium reaction vessel having a capacity of 500 mL together with 300 mL of an acidic chloride aqueous solution having a copper concentration of 60 g / L and a chlorine ion concentration of 200 g / L. The slurry concentration was 100 to 400 g / L.
Using an oil bath, the inside of the container is maintained at a temperature of 105 to 110 ° C., and chlorine gas is maintained so that the ORP is maintained at a predetermined constant value between 450 and 750 mV using an Ag / AgCl electrode as a reference electrode. The chlorine leaching process was performed by blowing. Sampling after 1, 3 and 6 hours from the start of treatment, analyzing the amount of existing substances in the leaching residue and copper ion concentration, iron ion concentration and sulfur concentration in the leachate, and leaching rate of copper and iron from copper concentrate And the sulfur oxidation rate was calculated.
さらに、塩素イオン濃度が２２〜４１９ｇ／Ｌになるように塩化銅、塩化鉄及び塩化ナトリウムを使用して調整して得た浸出始液を用いて、ＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）を５２０ｍＶに調整して浸出した。同時に浸出温度を１００〜１１０℃とした。 Furthermore, ORP (Ag / AgCl electrode standard) is set to 520 mV using a leaching start solution obtained by adjusting copper chloride, iron chloride and sodium chloride so that the chlorine ion concentration becomes 22 to 419 g / L. Adjusted and leached. At the same time, the leaching temperature was 100 to 110 ° C.
酸化還元電位（ＯＲＰ）と、銅と鉄の浸出率及び硫黄酸化率の関係を、表４に示し、スラリー濃度と、銅と鉄の浸出率の関係を図７に示す。さらに、図８に、終液塩素イオン濃度と、銅と鉄の浸出率及び硫黄酸化率の関係を示す。 Table 4 shows the relationship between the oxidation-reduction potential (ORP), the leaching rate of copper and iron, and the sulfur oxidation rate, and FIG. 7 shows the relationship between the slurry concentration and the leaching rate of copper and iron. Furthermore, FIG. 8 shows the relationship between the final solution chlorine ion concentration, the leaching rate of copper and iron, and the sulfur oxidation rate.
表４より、塩素浸出液のＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）を、好ましくは５００〜６００ｍＶ、より好ましくは５００〜５２０ｍＶに調整することによって、銅と鉄の高抽出率及び低硫黄酸化率が得られることが分る。また、図７より、浸出初期のスラリー濃度を、好ましくは１００〜４００ｇ／Ｌ、より好ましくは２５０〜４００ｇ／Ｌに調整することによって、銅と鉄の高抽出率及び低硫黄酸化率が得られることが分る。さらに、図８より、浸出終了時点での浸出終液に含まれる塩素イオン濃度を、好ましくは２００〜４００ｇ／Ｌ、より好ましくは２５０〜４００ｇ／Ｌに調整することによって、銅と鉄の高抽出率及び低硫黄酸化率が得られることが分る。 From Table 4, the high extraction rate and low sulfur oxidation rate of copper and iron can be obtained by adjusting the ORP (Ag / AgCl electrode standard) of the chlorine leaching solution to preferably 500 to 600 mV, more preferably 500 to 520 mV. I understand that. Moreover, from FIG. 7, the high extraction rate and low sulfur oxidation rate of copper and iron are obtained by adjusting the slurry concentration at the initial stage of leaching to preferably 100 to 400 g / L, more preferably 250 to 400 g / L. I understand that. Furthermore, from FIG. 8, the high extraction of copper and iron is achieved by adjusting the chlorine ion concentration contained in the leaching final solution at the end of leaching to preferably 200 to 400 g / L, more preferably 250 to 400 g / L. It can be seen that the rate and low sulfur oxidation rate are obtained.
（２）銅イオン還元処理工程
下記の浸出生成液Ａ、Ｂ、Ｃ、Ｄ、Ｅ、Ｆ、Ｇ、Ｈと、銅精鉱Ａ、Ｂ、Ｃ、Ｄ、Ｅ、Ｆを用いて、下記の浸出生成液の還元方法に従って、８９〜１０９℃の所定の還元温度で還元処理を行い、その際の最終のＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）を測定した。結果を表５に示す。なお、用いた浸出生成液の始液ＯＲＰ（Ａｇ／ＡｇＣｌ電極規準、９０℃）も参考に示す。
(2) Copper ion reduction treatment process Using the following leaching products A, B, C, D, E, F, G, H and copper concentrate A, B, C, D, E, F, the following According to the reduction method of the leaching product liquid, reduction treatment was performed at a predetermined reduction temperature of 89 to 109 ° C., and the final ORP (Ag / AgCl electrode standard) at that time was measured. The results are shown in Table 5. The starting liquid ORP (Ag / AgCl electrode standard, 90 ° C.) of the leaching product used is also shown for reference.
［浸出生成液］
（１）組成：いずれも、銅濃度３０ｇ／Ｌ、鉄濃度１００ｇ／Ｌ、塩化物イオン濃度２２０ｇ／Ｌである。
（２）始液ＯＲＰ（Ａｇ／ＡｇＣｌ電極規準、９０℃）
Ａ：５０８ｍＶ、Ｂ：４９０ｍＶ、Ｃ：４８１ｍＶ、Ｄ：４８２ｍＶ、Ｅ：４９０ｍＶ、Ｆ：４９１ｍＶ、Ｇ：４９５ｍＶ、Ｈ：４９８ｍＶ
[Leaching liquid]
(1) Composition: All have a copper concentration of 30 g / L, an iron concentration of 100 g / L, and a chloride ion concentration of 220 g / L.
(2) Start ORP (Ag / AgCl electrode standard, 90 ° C.)
A: 508 mV, B: 490 mV, C: 481 mV, D: 482 mV, E: 490 mV, F: 491 mV, G: 495 mV, H: 498 mV
［銅精鉱］
Ａ：銅精鉱（Ｄ５０：６０μｍ、化学組成：銅２６重量％、鉄２９重量％、硫黄２８重量％、その他１７重量％）。
Ｂ：銅精鉱Ａを、遊星ボールミル（セイシン企業製 プラネタリーミル ＳＫＦ−０４型）を用いて乾式粉砕して得た、Ｄ５０：３．７７μｍの粉砕物。なお、直径８ｍｍの鋼球を５０容量％充填した粉砕容器に入れて、１５０ｒｐｍで１時間処理した。
Ｃ：銅精鉱Ａを、ビーズミル（浅田鉄工製 ナノミル ＮＭ−Ｇ５Ｍ型）を用いて湿式粉砕して得た、Ｄ５０：１．０８μｍの粉砕物。なお、水を用いたスラリーの濃度を１００ｇ／Ｌとし、ビーズミルの粉砕室を３回パスさせた。
Ｄ：銅精鉱Ａを、ビーズミル（浅田鉄工製 ナノミル ＮＭ−Ｇ５Ｍ型）を用いて湿式粉砕して得た、Ｄ５０：１．３３μｍの粉砕物。なお、水を用いたスラリーの濃度を１００ｇ／Ｌとし、ビーズミルの粉砕室を２回パスさせた。
Ｅ：銅精鉱Ａを、ビーズミル（浅田鉄工製 ナノミル ＮＭ−Ｇ５Ｍ型）を用いて湿式粉砕して得た、Ｄ５０：２．２１μｍの粉砕物。なお、水を用いたスラリーの濃度を１００ｇ／Ｌとし、ビーズミルの粉砕室を１回パスさせた。
Ｆ：銅精鉱Ａを、ビーズミル（浅田鉄工製 ナノミル ＮＭ−Ｇ５Ｍ型）を用いて湿式粉砕して得た、Ｄ５０：０．７６μｍの粉砕物。なお、水を用いたスラリーの濃度を１００ｇ／Ｌとし、ビーズミルの粉砕室を１時間循環させた。
[Copper concentrate]
A: Copper concentrate (D50: 60 μm, chemical composition: copper 26% by weight, iron 29% by weight, sulfur 28% by weight, other 17% by weight).
B: D50: 3.77 μm pulverized product obtained by dry pulverizing copper concentrate A using a planetary ball mill (planetary mill SKF-04, manufactured by Seishin Enterprise). In addition, it put into the grinding | pulverization container filled with 50 volume% of steel balls with a diameter of 8 mm, and processed at 150 rpm for 1 hour.
C: D50: 1.08 μm pulverized product obtained by wet-pulverizing copper concentrate A using a bead mill (Nanomill NM-G5M type manufactured by Asada Tekko). In addition, the density | concentration of the slurry using water was 100 g / L, and the grinding | pulverization chamber of the bead mill was passed 3 times.
D: D50: 1.33 μm pulverized product obtained by wet-pulverizing copper concentrate A using a bead mill (Nanomill NM-G5M type manufactured by Asada Tekko). In addition, the density | concentration of the slurry using water was 100 g / L, and the grinding | pulverization chamber of the bead mill was passed twice.
E: A pulverized product of D50: 2.21 μm obtained by wet-grinding copper concentrate A using a bead mill (Nanomill NM-G5M type manufactured by Asada Tekko). In addition, the density | concentration of the slurry using water was 100 g / L, and the grinding | pulverization chamber of the bead mill was passed once.
F: D50: 0.76 μm pulverized product obtained by wet pulverizing copper concentrate A using a bead mill (Nanomill NM-G5M type manufactured by Asada Tekko). In addition, the density | concentration of the slurry using water was 100 g / L, and the grinding | pulverization chamber of the bead mill was circulated for 1 hour.
［浸出生成液の還元方法］
所定組成の浸出生成液に所定の平均粒子径（Ｄ５０）に調整された銅精鉱５０ｇを、初期スラリー濃度が１００ｇ／Ｌになるように５００ｍＬのガラスビーカーに装入して撹拌し、所定の還元温度に加熱して３時間保持した。
[Reduction method of leaching product]
50 g of copper concentrate adjusted to a predetermined average particle size (D50) in a leaching product solution having a predetermined composition was charged into a 500 mL glass beaker so that the initial slurry concentration would be 100 g / L, and stirred. Heated to the reduction temperature and held for 3 hours.
表５より、用いた銅精鉱のＤ５０から上記の式１を満足する温度（Ａ）を求め、前記浸出生成液の還元温度を９０〜１１０℃の範囲でかつ温度（Ａ）以上に調整して行った場合（Ｎｏ．１〜７）には、４００ｍＶ以下のＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）が得られ、特に式２を満足する温度（Ａ）以上に調整して行った場合（Ｎｏ．１）には、３８０ｍＶ以下が得られることが分かる。これに対して、Ｄ５０と還元温度の関係が上記式１又は２の条件に合わない場合（Ｎｏ．８〜１２）には、４００ｍＶ以下のＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）が得られないことが分かる。したがって、銅精鉱のＤ５０と還元温度の関係を、上記式１又は２を満足する条件に調整することが好ましい。
From Table 5, the temperature (A) satisfying the above-mentioned
（３）溶媒抽出工程
還元生成液Ａ、Ｂ、Ｃ、Ｄを用いて、これら還元生成液を溶媒抽出に付し、銅を抽出した後、逆抽出して、第１銅イオンを含む逆抽出生成液と第１鉄イオンを含む抽出残液とを得て、評価した。
(3) Solvent extraction step Using the reduction product liquids A, B, C, and D, these reduction product liquids are subjected to solvent extraction to extract copper, followed by back extraction and back extraction containing the first copper ions. A product solution and an extraction residue containing ferrous ions were obtained and evaluated.
還元生成液Ａとして、Ｃｕ７５ｇ／Ｌ、Ｆｅ５１ｇ／Ｌの濃度を有する上記工程で得られた還元生成液を用い、この５００ｍＬに銅粉を加えて酸化還元電位（Ａｇ／ＡｇＣｌ電極規準）を３００ｍＶに調整し、還元生成液中の銅イオンを完全に一価に還元した。この液を溶媒抽出の始液とし、希釈剤（商品名シェルゾールＡ、昭和シェル石油（株）製）で４０容量％の濃度に調整したトリブチルフォスフェイト（商品名ＴＢＰ、大八化学（株）製）５００ｍＬと混合して、液温を５０℃として、１０分間振とうして溶媒抽出を行った。次いで、有機溶媒相と水相を分離し、この有機溶媒相に、塩酸でｐＨ１に調整した２５０ｍＬの水を混合し１０分間振とうして、逆抽出した。
得られた逆抽出生成液の組成を、溶媒抽出の始液及び抽出残液とともに表６に示す。
As the reduction product A, the reduction product obtained in the above step having a concentration of Cu 75 g / L and Fe 51 g / L was used, and copper powder was added to 500 mL to bring the oxidation-reduction potential (Ag / AgCl electrode standard) to 300 mV. The copper ions in the reduction product solution were completely reduced to a monovalent state. Tributyl phosphate (trade name TBP, Daihachi Chemical Co., Ltd.) adjusted to a concentration of 40% by volume with a diluent (trade name Shellzol A, manufactured by Showa Shell Sekiyu K.K.) was used as the starting solution for solvent extraction. (Made) 500 mL was mixed, the temperature of the solution was 50 ° C., and the mixture was shaken for 10 minutes for solvent extraction. Next, the organic solvent phase and the aqueous phase were separated, and 250 mL of water adjusted to
The composition of the obtained back extraction product liquid is shown in Table 6 together with the solvent extraction start liquid and the extraction residual liquid.
表６より、抽出剤として、トリブチルフォスフェイトを用いることによって、銅が鉄に対して、選択的にかつ高効率で分離されることが分る。 It can be seen from Table 6 that copper is selectively and efficiently separated from iron by using tributyl phosphate as an extractant.
還元生成液Ｂとして、Ｃｕ８０ｇ／Ｌ、Ｆｅ５０ｇ／Ｌの濃度の塩化第２銅と塩化第２鉄からなる合成水溶液に、食塩を添加して塩素イオン濃度を２００ｇ／Ｌとし、さらにこの液を６０℃に加温し、鉄粉を投入してＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）を３００ｍＶに調整した液を用いた。前記液を、室温にてトリブチルフォスフェイト濃度が４０〜１００容積％の所定値にケロシンで希釈して得た有機溶媒と接触混合させ、トリブチルフォスフェイト中に金属イオンを抽出し、銅、鉄それぞれの抽出率を求めた。なお、接触混合時には混合時の気液界面接触による酸化を防ぐため窒素により不活性雰囲気とした。結果を図９、図１０に示す。 As reduction product liquid B, sodium chloride was added to a synthetic aqueous solution composed of cupric chloride and ferric chloride at a concentration of Cu 80 g / L and Fe 50 g / L to make the chlorine ion concentration 200 g / L. The solution was heated to 0 ° C., charged with iron powder, and adjusted to ORP (Ag / AgCl electrode standard) of 300 mV. The liquid is contact-mixed with an organic solvent obtained by diluting with kerosene to a predetermined value of tributyl phosphate concentration of 40 to 100% by volume at room temperature, metal ions are extracted into tributyl phosphate, and each of copper and iron is extracted. The extraction rate was determined. During contact mixing, an inert atmosphere was established with nitrogen to prevent oxidation due to gas-liquid interface contact during mixing. The results are shown in FIGS.
図９より、トリブチルフォスフェイトの濃度を上昇させると銅の抽出率は大きくなるが、鉄の抽出率はあまり変化がないことが分る。図１０より、トリブチルフォスフェイトの濃度が高いと銅／鉄の分離係数も良くなることが分る。 From FIG. 9, it can be seen that when the concentration of tributyl phosphate is increased, the extraction rate of copper increases, but the extraction rate of iron does not change much. From FIG. 10, it can be seen that the higher the concentration of tributyl phosphate, the better the copper / iron separation factor.
還元生成液Ｃとして、Ｃｕ１１８ｇ／Ｌ、Ｆｅ９０ｇ／Ｌ、Ａｇ９ｍｇ／Ｌの濃度となるように塩化第２銅、塩化第２鉄及び塩化銀を含む銅精鉱浸出液を作製し、さらにこの液を６０℃に加温し、鉄粉を投入してＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）を３００ｍＶに調整した液を用いた。前記液を、室温にてトリブチルフォスフェイトと接触混合させ、トリブチルフォスフェイト中に金属イオンを抽出した。なお、接触混合時には混合時の気液界面接触による酸化を防ぐため窒素により不活性雰囲気とした。次に、塩酸でｐＨ１．０に調整し、食塩で塩素イオン濃度を５０ｇ／Ｌに調整した逆抽出液を用いて、５０℃で、トリブチルフォスフェイト中の銅イオンを逆抽出した。その後、抽出後の抽出残液と逆抽出生成液のＣｕ、Ｆｅ、Ａｇを分析した。
As reduction product liquid C, a copper concentrate leachate containing cupric chloride, ferric chloride and silver chloride so as to have a concentration of Cu 118 g / L, Fe 90 g / L, and
その結果、抽出残液中の濃度はＣｕ６０ｇ／Ｌ、Ｆｅ９０ｇ／Ｌ、Ａｇ８ｍｇ／Ｌであり、逆抽出生成液中の濃度はＣｕ２８ｇ／Ｌ、Ｆｅ５ｇ／Ｌ、Ａｇ１ｍｇ／Ｌ以下であった。これより、銀はトリブチルフォスフェイトに抽出されず、抽出残液に残り、銅と銀は殆ど完全に分離できることが分る。
As a result, the concentrations in the extraction residual liquid were Cu 60 g / L, Fe 90 g / L, and
還元生成液Ｄとして、Ｃｕ５０ｇ／Ｌ、Ｆｅ７０ｇ／Ｌの濃度の塩化第２銅、塩化第２鉄からなる合成水溶液に、食塩を添加して塩素イオン濃度を２００ｇ／Ｌとし、さらにこの液を６０℃に加温し、鉄粉を投入してＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）を２７５ｍＶに調整した液を用いた。前記液を、室温にてトリブチルフォスフェイトと接触混合させ、トリブチルフォスフェイト中に金属イオンを抽出した。なお、接触混合時には混合時の気液界面接触による酸化を防ぐため窒素により不活性雰囲気とした。次に、塩酸でｐＨ０．５に、食塩で塩素イオン濃度を１００ｇ／Ｌに調整した逆抽出液を用いて、３０℃、４０℃、６０℃、７５℃で、トリブチルフォスフェイト中の銅を逆抽出した。その後、Ｃｕを分析して逆抽出率を求めた。結果を図１１に示す。 As reduction product liquid D, sodium chloride was added to a synthetic aqueous solution consisting of cupric chloride and ferric chloride at a concentration of Cu 50 g / L and Fe 70 g / L to make the chloride ion concentration 200 g / L. The solution was heated to 0 ° C., iron powder was added, and ORP (Ag / AgCl electrode standard) was adjusted to 275 mV. The liquid was contact-mixed with tributyl phosphate at room temperature, and metal ions were extracted into tributyl phosphate. During contact mixing, an inert atmosphere was established with nitrogen to prevent oxidation due to gas-liquid interface contact during mixing. Next, reverse the copper in tributyl phosphate at 30 ° C, 40 ° C, 60 ° C and 75 ° C using a back extract adjusted to pH 0.5 with hydrochloric acid and the chloride ion concentration to 100 g / L with sodium chloride. Extracted. Then, Cu was analyzed and the back extraction rate was calculated | required. The results are shown in FIG.
図１１より、温度を上げることにより逆抽出率は大きくなり、トリブチルフォスフェイト中の銅の５０％以上を逆抽出するには、５０℃以上が好ましいことが分る。 From FIG. 11, it can be seen that the back extraction rate increases with increasing temperature, and 50 ° C. or higher is preferable for back extraction of 50% or more of copper in tributyl phosphate.
（４）銅電解採取工程
逆抽出生成液を、以下の隔膜電解槽を用いて電解採取に付し、電着銅と電解尾液を得て、評価した。
隔膜電解槽の構造は、図６に示すものである。ここで、陽極は、７０ｍｍ×７０ｍｍの不溶性電極（商品名ＤＳＥ、ペルメレック電極（株）製）で、また陰極は、６０ｍｍ×６５ｍｍの純銅板である。陰極室と陽極室の間を、目が細かい通水度の低いテトロン製の濾布（通水度０．０６Ｌ／ｍ２．ｓ）を隔膜として用いて分離した。陰極室及び陽極室のそれぞれにオーバーフロー方式の排液口を設け、それぞれ別に排液できるようにした。そのとき、陰極室のオーバーフローレベルは陽極室のそれよりも若干高めとし、液面差を付けてその圧力で陽極室の塩素ガスや塩素ガスを含んだ液が陰極室に入り込まないようにした。
(4) Copper electrowinning process The back-extraction product solution was subjected to electrowinning using the following diaphragm electrolytic cell to obtain electrodeposited copper and an electrolytic tail solution, and evaluated.
The structure of the diaphragm electrolytic cell is shown in FIG. Here, the anode is a 70 mm × 70 mm insoluble electrode (trade name DSE, manufactured by Permerek Electrode Co., Ltd.), and the cathode is a pure copper plate of 60 mm × 65 mm. The cathode chamber and the anode chamber were separated by using a fine Tetron filter cloth (water permeability 0.06 L / m 2 · s) as a diaphragm. Each of the cathode chamber and the anode chamber was provided with an overflow type drainage port so that it could be drained separately. At that time, the overflow level of the cathode chamber was slightly higher than that of the anode chamber, and a liquid level difference was added so that chlorine gas in the anode chamber or a liquid containing chlorine gas did not enter the cathode chamber at that pressure.
この電解槽の陰極室に、銅濃度８０ｇ／Ｌの塩化第１銅水溶液を２．５ｍＬ／ｍｉｎの給液量で供給し、一方陽極室には鉄濃度８０ｇ／Ｌの塩化第２鉄水溶液を供給した。このとき、陰極室に供給する塩化第１銅水溶液のＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）は３０９ｍＶであった。 A cuprous chloride aqueous solution having a copper concentration of 80 g / L is supplied to the cathode chamber of the electrolytic cell at a supply amount of 2.5 mL / min, while a ferric chloride aqueous solution having an iron concentration of 80 g / L is supplied to the anode chamber. Supplied. At this time, the ORP (Ag / AgCl electrode standard) of the cuprous chloride aqueous solution supplied to the cathode chamber was 309 mV.
電流を約１．２Ａ流し、電流密度３０８Ａ／ｍ２で銅電解採取を行った。このとき、電着した銅からの電流効率は９３％であり、一般的な塩化第１銅の銅電解と同等の電流効率（８５〜９５％）が得られることが分る。また、１０分間の陰極室への給液２５．５ｍＬに対して、陰極室からの排液量は２５．２ｍＬであり、陰極室液の回収率は９９％となった。即ち、陰極室液が陽極室へ少量移動し、陽極室液の陰極室への流入は無いと判断できる。このときの陰極室廃液のＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）は３０７ｍＶであり、電着による銅イオンの消費以外は陰極室液の性状は殆ど変化がない。これより、陰極に供給した液の殆どを銅イオン量が減っただけの状態で回収することができ、陰極廃液は溶媒抽出の逆抽出液として好適である。 Copper electrowinning was performed at a current density of 308 A / m 2 at a current of about 1.2 A. At this time, the current efficiency from the electrodeposited copper is 93%, and it is understood that the current efficiency (85 to 95%) equivalent to that of general copper chloride electrolysis is obtained. Moreover, with respect to 25.5 mL of liquid supply to the cathode chamber for 10 minutes, the amount of drainage from the cathode chamber was 25.2 mL, and the recovery rate of the cathode chamber liquid was 99%. That is, it can be determined that the cathodic chamber liquid moves a small amount to the anode chamber and that the anode chamber liquid does not flow into the cathode chamber. The ORP (Ag / AgCl electrode standard) of the cathode chamber waste liquid at this time is 307 mV, and there is almost no change in the properties of the cathode chamber liquid other than the consumption of copper ions by electrodeposition. As a result, most of the liquid supplied to the cathode can be recovered with a reduced amount of copper ions, and the cathode waste liquid is suitable as a back extraction liquid for solvent extraction.
また、陽極給液である塩化第２鉄水溶液のＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）は６００ｍＶであったが、陽極での酸化反応によって塩素ガスが発生し、陽極廃液のＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）は９００ｍＶ以上の強酸化液となった。これにより、陽極廃液と塩素ガスとを塩素浸出工程に循環することによって、より高い酸化力が得られる。 Further, the ORP (Ag / AgCl electrode standard) of the ferric chloride aqueous solution as the anode supply liquid was 600 mV, but chlorine gas was generated by the oxidation reaction at the anode, and the ORP (Ag / AgCl electrode standard of the anode waste liquid). ) Became a strong oxidizing solution of 900 mV or higher. Thereby, a higher oxidizing power can be obtained by circulating the anode waste liquid and the chlorine gas to the chlorine leaching step.
また、上記の銅電解採取工程で得られた電着銅の分析値の一例を、表７に示す。表７より、純度９９．９９重量％の不純物が少ない電着銅が得られることが分る。 Table 7 shows an example of the analytical value of the electrodeposited copper obtained in the above copper electrowinning process. From Table 7, it can be seen that an electrodeposited copper with a purity of 99.99% by weight and less impurities can be obtained.
（５）浄液工程
溶媒抽出工程において、抽出残液には鉄の他に、２〜５ｇ／Ｌの濃度の銅、０．５ｇ／Ｌ程度の濃度のヒ素、アンチモン、ニッケル、亜鉛、鉛、及びカドミウム、さらに０．１ｇ／Ｌ程度の濃度の銀等多くの随伴する有価金属が含まれており、鉄回収にとっては不純物となる。
(5) Liquid purification step In the solvent extraction step, the extraction residual liquid contains, in addition to iron, copper having a concentration of 2 to 5 g / L, arsenic having a concentration of about 0.5 g / L, antimony, nickel, zinc, lead, In addition, cadmium and many accompanying valuable metals such as silver having a concentration of about 0.1 g / L are contained, which becomes an impurity for iron recovery.
これらの浄液処理方法として、硫化処理、セメンテーション処理、又は中和処理を用いた。抽出残液を５０℃に加温し、各方法で１時間処理した。中和処理は、ｐＨ３になるように消石灰を添加して中和した。また、硫化処理は、ｐＨ０．５で水硫化ナトリウムの水溶液を不純物を硫化するのに必要な当量の２倍ないし５倍を添加した。セメンテーション処理は、鉄粉を添加してＯＲＰを調整し、ｐＨ１．４でＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）−７１ｍＶ、ｐＨ１．１でＯＲＰ（Ａｇ／ＡｇＣｌ電極規準）−２５０ｍＶで処理した。各浄液処理方法での各元素の除去率の結果を、表８に示す。 As these liquid treatment methods, sulfuration treatment, cementation treatment, or neutralization treatment was used. The extraction residue was heated to 50 ° C. and treated for 1 hour by each method. Neutralization treatment was neutralized by adding slaked lime so that the pH was 3. In addition, in the sulfiding treatment, an aqueous solution of sodium hydrosulfide having a pH of 0.5 was added 2 to 5 times the equivalent amount required to sulfidize impurities. In the cementation treatment, iron powder was added to adjust the ORP, and the treatment was performed at ORP (Ag / AgCl electrode standard) -71 mV at pH 1.4 and ORP (Ag / AgCl electrode standard) -250 mV at pH 1.1. Table 8 shows the results of the removal rate of each element in each liquid treatment method.
表８より、各浄液処理方法で不純物を効果的に除去できること分る。例えば、ヒ素又はアンチモンは中和で除去することができ、一方、鉛又はニッケルを除去するには硫化反応が適している。また、銀を回収するにはセメンテーション又は硫化が効果的である。なお、これらの処理は、銅精鉱に含まれる不純物に応じて使い分ければよいので、必ずしもすべての処理が必要であると言うことではない。
また、各処理方法を適宜使い分けることで不純物を選択的に分離し、有効利用又は処理に適した形態で回収することも可能である。
From Table 8, it can be seen that impurities can be effectively removed by each liquid treatment method. For example, arsenic or antimony can be removed by neutralization, while a sulfurization reaction is suitable for removing lead or nickel. Also, cementation or sulfidation is effective for recovering silver. In addition, since these processes should just be properly used according to the impurity contained in copper concentrate, it does not necessarily mean that all the processes are required.
Moreover, it is also possible to selectively separate impurities by properly using each processing method, and collect them in a form suitable for effective use or processing.
（６）鉄電解採取工程
浄液処理した鉄を含む精製液を用いて、鉄電解採取を行い、評価した。
電解槽は、容量５００ｍＬで、その内部に電極面積が６０ｍｍ×６０ｍｍの不溶性陽極（商品名ＤＳＥ、ペルメレック電極（株）製）１枚と、同じ面積になるようにマスキングした陰極（チタン板）２枚を距離が６０ｍｍになるように設置した。
(6) Iron electrowinning step Iron electrowinning was performed and evaluated using a purified solution containing purified iron.
The electrolytic cell has a capacity of 500 mL, and an insoluble anode (trade name: DSE, manufactured by Permerek Electrode Co., Ltd.) having an electrode area of 60 mm × 60 mm inside, and a cathode (titanium plate) 2 masked to have the same area. The sheets were placed so that the distance was 60 mm.
電解液を毎分１ｍＬあるいは毎分５ｍＬの割合で、濾布で仕切られた陰極側にポンプで給液しながら、電流密度が２００Ａ／ｍ２となるように通電して電解採取した。なお、給液した電解液は繰り返して使用せず、そのまま予備槽に貯留した。この場合、理論電着量は毎時１．５ｇとなるので、給液中の鉄イオンは、毎分１ｍＬで給液したときは、毎時１．５ｇと理論電着量と同じで、毎分５ｍＬで給液したときは、毎時７．５ｇと理論電着量の５倍供給されることになる。 While supplying the electrolytic solution at a rate of 1 mL per minute or 5 mL per minute with a pump to the cathode side partitioned by the filter cloth, the current was electrified so that the current density was 200 A / m 2 . In addition, the supplied electrolyte solution was not used repeatedly, but was stored in a preliminary tank as it was. In this case, since the theoretical electrodeposition amount is 1.5 g / h, when the iron ions in the liquid supply are supplied at 1 mL / min, the theoretical electrodeposition amount is 1.5 g / h, which is 5 mL / min. When the liquid is supplied, 7.5 g per hour is supplied, which is 5 times the theoretical electrodeposition amount.
５時間の通電後、カソード表面を洗浄して電着量を測定し電流効率を算出した。電流効率は、９７％であった。
毎分１ｍＬで給液した場合には、浴電圧は２．９Ｖとなって、陽極から塩素ガスの発生が目視で確認された。一方、毎分５ｍＬで給液すると陽極からの塩素ガスの発生は観察されず、浴電圧も２．３Ｖまで低下した。即ち、本発明の電解方法によって、電力コストを低減できることが分る。
After energization for 5 hours, the cathode surface was washed, the amount of electrodeposition was measured, and the current efficiency was calculated. The current efficiency was 97%.
When the liquid was supplied at 1 mL per minute, the bath voltage was 2.9 V, and the generation of chlorine gas from the anode was visually confirmed. On the other hand, when the liquid was supplied at 5 mL / min, generation of chlorine gas from the anode was not observed, and the bath voltage was lowered to 2.3V. That is, it can be seen that the power cost can be reduced by the electrolysis method of the present invention.
（７）浸出残渣処理工程
塩素浸出工程の浸出残渣から、元素状硫黄を分離し、貴金属を含む残滓を回収する蒸留を行った。塩素浸出工程で得られた浸出残渣から１５０ｇを分取し、これを石英製のボートに入れて管状炉中に挿入した。炉内には窒素ガスを流し、３２０℃になるように加熱した。ガスの流れ出る端部に冷却管を設け空冷した。加熱開始から４時間経過後にガスを流したまま電源を切り、炉内温度が概ね７０℃以下となってから残渣と硫黄を取り出した。表９に、浸出残渣と脱硫黄後の残渣の化学分析値、及び硫黄の形態分析結果を示す。
(7) Leaching residue treatment process The elemental sulfur was separated from the leaching residue of the chlorine leaching process, and distillation was performed to recover the residue containing noble metals. From the leaching residue obtained in the chlorine leaching step, 150 g was collected, put into a quartz boat, and inserted into a tubular furnace. Nitrogen gas was flowed into the furnace and heated to 320 ° C. A cooling pipe was provided at the end where the gas flowed out to cool it down. After 4 hours from the start of heating, the power was turned off with the gas flowing, and the residue and sulfur were taken out after the furnace temperature became approximately 70 ° C. or lower. Table 9 shows the chemical analysis values of the leaching residue and the residue after desulfurization, and the results of the sulfur form analysis.
表９より、浸出残渣から硫黄が除去され、その分だけ金が濃縮されていることが分る。 From Table 9, it can be seen that sulfur is removed from the leaching residue and gold is concentrated accordingly.
以上より明らかなように、本発明の硫化銅鉱物を含む銅原料の精錬方法は、硫化銅鉱物を含む銅原料の湿式精錬法として利用され、原料中に含まれる硫黄の酸化を抑制しながら、銅を浸出して一価銅電解で回収する方法として、また同時に随伴する有価金属も回収し有効に活用する方法として、さらに浸出残渣などの廃棄物を可能な限り減少する方法等として有用であり、特に難浸出性の黄銅鉱の精錬に用いるのに適している。 As is clear from the above, the method for refining a copper raw material containing a copper sulfide mineral of the present invention is used as a wet refining method for a copper raw material containing a copper sulfide mineral, while suppressing the oxidation of sulfur contained in the raw material, It is useful as a method for leaching copper and recovering it by monovalent copper electrolysis, as well as a method for recovering and effectively utilizing the accompanying valuable metals, and a method for reducing waste such as leaching residue as much as possible. In particular, it is suitable for use in refining hardly leachable chalcopyrite.
１ 塩素浸出工程
２ 銅イオン還元処理工程
３ 溶媒抽出工程
４ 銅電解採取工程
５ 浄液工程
６ 鉄電解採取工程
７ 浸出残渣処理工程
８ 銅原料
９ 電着銅
１０ 電着鉄
１１ 陰極室
１２ 陽極室
１３ 隔膜
１４ 陰極
１５ 陽極
１６ 陰極オーバーフロー口
１７ 陽極オーバーフロー口
１８ 陰極給液
１９ 陽極給液
２０ 陰極室液
２１ 陽極室液
２２ 陰極廃液
２３ 陽極廃液
DESCRIPTION OF
Claims (22)
（１）前記銅原料を酸性塩化物水溶液中で塩素による浸出に付し、該液中に銅を溶出させ、銅イオンを含む浸出生成液と元素状硫黄を含む残渣とを形成する塩素浸出工程、
（２）前記浸出生成液に還元剤を添加し、銅イオンを還元して第１銅イオンを含む還元生成液を得る銅イオン還元処理工程、
（３）前記還元生成液を溶媒抽出に付し、銅を抽出した後、逆抽出して、第１銅イオンを含む逆抽出生成液と第１鉄イオンを含む抽出残液とを得る溶媒抽出工程、
（４）前記逆抽出生成液を電解採取に付し、電着銅と電解尾液とを形成する銅電解採取工程、及び
（５）前記抽出残液を電解採取に付し、電着鉄と塩化鉄水溶液からなる鉄電解尾液とを形成する鉄電解採取工程、を含むことを特徴とする硫化銅鉱物を含む銅原料の精錬方法。 In a method of recovering valuable metals associated with copper by a copper refining method from a copper raw material containing copper sulfide minerals,
(1) Chlorine leaching process in which the copper raw material is subjected to leaching with chlorine in an aqueous acid chloride solution, and copper is eluted in the liquid to form a leaching product liquid containing copper ions and a residue containing elemental sulfur. ,
(2) A copper ion reduction treatment step of adding a reducing agent to the leaching product liquid and reducing the copper ions to obtain a reduced product liquid containing cuprous ions,
(3) Solvent extraction by subjecting the reduction product solution to solvent extraction to extract copper and back-extracting to obtain a back-extraction product solution containing first copper ions and an extraction residual solution containing ferrous ions. Process,
(4) subjecting the back-extraction product solution to electrowinning to form a copper electrowinning step for forming electrodeposited copper and electrolytic tail liquor; and (5) subjecting the extraction residual solution to electrowinning, A method for refining a copper raw material containing a copper sulfide mineral, comprising: an iron electrowinning step of forming an iron electrolytic tail solution comprising an aqueous iron chloride solution.
式１：Ａ（℃）＝６．７９×Ｌｎ（Ｂ）＋８１．５
（式中、Ｂは、銅精鉱の体積頻度累積が５０容量％に相当する平均粒子径（Ｄ５０）を意味し、マイクロトラックを用いて測定された単位μｍの数値である。） While using copper concentrate whose main mineral is chalcopyrite prepared to have an average particle size (D50) of 0.5 to 60 μm as the copper sulfide mineral, the reduction temperature of the leaching product liquid is 90 to 110 ° C. The method for refining a copper raw material containing a copper sulfide mineral according to claim 6, wherein the temperature is adjusted to a temperature (A) calculated from the following formula 1 or more.
Formula 1: A (° C.) = 6.79 × Ln (B) +81.5
(In the formula, B means an average particle diameter (D50) in which the volume frequency accumulation of copper concentrate corresponds to 50% by volume, and is a numerical value of unit μm measured using a microtrack.)
式２：Ａ（℃）＝７．０４×Ｌｎ（Ｂ）＋９５．２
（式中、Ｂは、銅精鉱の体積頻度累積が５０容量％に相当する平均粒子径（Ｄ５０）を意味し、マイクロトラックを用いて測定された単位μｍの数値である。） While using copper concentrate whose main mineral is chalcopyrite prepared to have an average particle size (D50) of 0.5 to 60 μm as the copper sulfide mineral, the reduction temperature of the leaching product liquid is 90 to 110 ° C. The method for refining a copper raw material containing a copper sulfide mineral according to claim 6, wherein the temperature is adjusted to a temperature (A) calculated by the following formula 2 within a range of:
Formula 2: A (° C.) = 7.04 × Ln (B) +95.2
(In the formula, B means an average particle diameter (D50) in which the volume frequency accumulation of copper concentrate corresponds to 50% by volume, and is a numerical value of unit μm measured using a microtrack.)
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| JP2003315124A JP4352823B2 (en) | 2002-11-18 | 2003-09-08 | Method for refining copper raw materials containing copper sulfide minerals |
| AU2003261548A AU2003261548A1 (en) | 2002-11-18 | 2003-11-10 | Process for refining raw copper material containing copper sulfide mineral |
| US10/706,276 US20040144208A1 (en) | 2002-11-18 | 2003-11-13 | Process for refining raw copper material containing copper sulfide mineral |
| CNB2003101153044A CN1262677C (en) | 2002-11-18 | 2003-11-14 | Refined method for copper materials containing copper sulphide minerals |
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| JP2003282248 | 2003-07-30 | ||
| JP2003315124A JP4352823B2 (en) | 2002-11-18 | 2003-09-08 | Method for refining copper raw materials containing copper sulfide minerals |
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| CN114502752B (en) * | 2019-07-11 | 2023-10-13 | 公共股份公司矿业和冶金公司诺里尔斯克镍 | Method for processing copper sulfide and nickel sulfide materials |
Also Published As
| Publication number | Publication date |
|---|---|
| CN1262677C (en) | 2006-07-05 |
| CN1500892A (en) | 2004-06-02 |
| US20040144208A1 (en) | 2004-07-29 |
| AU2003261548A1 (en) | 2004-06-03 |
| JP4352823B2 (en) | 2009-10-28 |
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