WO2013085052A1 - 希土類元素の回収方法 - Google Patents
希土類元素の回収方法 Download PDFInfo
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- WO2013085052A1 WO2013085052A1 PCT/JP2012/081856 JP2012081856W WO2013085052A1 WO 2013085052 A1 WO2013085052 A1 WO 2013085052A1 JP 2012081856 W JP2012081856 W JP 2012081856W WO 2013085052 A1 WO2013085052 A1 WO 2013085052A1
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- WIPO (PCT)
- Prior art keywords
- rare earth
- leaching
- earth elements
- earth element
- recovering
- Prior art date
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- 229910052761 rare earth metal Inorganic materials 0.000 title claims abstract description 235
- 238000000034 method Methods 0.000 title claims abstract description 146
- 238000011084 recovery Methods 0.000 title claims abstract description 67
- 238000002386 leaching Methods 0.000 claims abstract description 127
- 239000007788 liquid Substances 0.000 claims abstract description 106
- 238000000605 extraction Methods 0.000 claims abstract description 86
- 238000000926 separation method Methods 0.000 claims abstract description 73
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 claims abstract description 40
- 239000002994 raw material Substances 0.000 claims abstract description 34
- 230000008569 process Effects 0.000 claims abstract description 31
- 239000007787 solid Substances 0.000 claims abstract description 26
- 239000002253 acid Substances 0.000 claims abstract description 25
- 239000002002 slurry Substances 0.000 claims abstract description 22
- 229910052719 titanium Inorganic materials 0.000 claims abstract description 17
- GRYLNZFGIOXLOG-UHFFFAOYSA-N Nitric acid Chemical compound O[N+]([O-])=O GRYLNZFGIOXLOG-UHFFFAOYSA-N 0.000 claims abstract description 6
- 229910017604 nitric acid Inorganic materials 0.000 claims abstract description 6
- 229910001570 bauxite Inorganic materials 0.000 claims description 82
- 238000000638 solvent extraction Methods 0.000 claims description 65
- 239000007864 aqueous solution Substances 0.000 claims description 49
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 42
- 239000000243 solution Substances 0.000 claims description 35
- 229910052782 aluminium Inorganic materials 0.000 claims description 32
- MUBZPKHOEPUJKR-UHFFFAOYSA-N Oxalic acid Chemical compound OC(=O)C(O)=O MUBZPKHOEPUJKR-UHFFFAOYSA-N 0.000 claims description 27
- VLKZOEOYAKHREP-UHFFFAOYSA-N n-Hexane Chemical compound CCCCCC VLKZOEOYAKHREP-UHFFFAOYSA-N 0.000 claims description 27
- 229910052742 iron Inorganic materials 0.000 claims description 24
- 150000004679 hydroxides Chemical class 0.000 claims description 21
- 239000003350 kerosene Substances 0.000 claims description 20
- -1 octanol Chemical compound 0.000 claims description 20
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims description 17
- 230000029087 digestion Effects 0.000 claims description 16
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 15
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 claims description 15
- 239000000203 mixture Substances 0.000 claims description 15
- 239000000047 product Substances 0.000 claims description 15
- 150000002500 ions Chemical class 0.000 claims description 14
- UHOVQNZJYSORNB-UHFFFAOYSA-N Benzene Chemical compound C1=CC=CC=C1 UHOVQNZJYSORNB-UHFFFAOYSA-N 0.000 claims description 12
- YXFVVABEGXRONW-UHFFFAOYSA-N Toluene Chemical compound CC1=CC=CC=C1 YXFVVABEGXRONW-UHFFFAOYSA-N 0.000 claims description 12
- 238000007865 diluting Methods 0.000 claims description 11
- 239000007800 oxidant agent Substances 0.000 claims description 11
- 239000002244 precipitate Substances 0.000 claims description 11
- 150000002148 esters Chemical class 0.000 claims description 10
- VLTRZXGMWDSKGL-UHFFFAOYSA-N perchloric acid Chemical compound OCl(=O)(=O)=O VLTRZXGMWDSKGL-UHFFFAOYSA-N 0.000 claims description 10
- 239000000839 emulsion Substances 0.000 claims description 9
- 238000010979 pH adjustment Methods 0.000 claims description 9
- 239000003002 pH adjusting agent Substances 0.000 claims description 8
- MHAJPDPJQMAIIY-UHFFFAOYSA-N Hydrogen peroxide Chemical group OO MHAJPDPJQMAIIY-UHFFFAOYSA-N 0.000 claims description 7
- 239000011260 aqueous acid Substances 0.000 claims description 7
- 238000001914 filtration Methods 0.000 claims description 6
- 238000010438 heat treatment Methods 0.000 claims description 6
- 235000006408 oxalic acid Nutrition 0.000 claims description 6
- 239000000126 substance Substances 0.000 claims description 6
- STCOOQWBFONSKY-UHFFFAOYSA-N tributyl phosphate Chemical compound CCCCOP(=O)(OCCCC)OCCCC STCOOQWBFONSKY-UHFFFAOYSA-N 0.000 claims description 6
- QAOWNCQODCNURD-UHFFFAOYSA-L Sulfate Chemical compound [O-]S([O-])(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-L 0.000 claims description 5
- 229910010413 TiO 2 Inorganic materials 0.000 claims description 5
- 239000000463 material Substances 0.000 claims description 5
- KBPLFHHGFOOTCA-UHFFFAOYSA-N 1-Octanol Chemical compound CCCCCCCCO KBPLFHHGFOOTCA-UHFFFAOYSA-N 0.000 claims description 4
- NBIIXXVUZAFLBC-UHFFFAOYSA-N Phosphoric acid Chemical compound OP(O)(O)=O NBIIXXVUZAFLBC-UHFFFAOYSA-N 0.000 claims description 4
- 150000004945 aromatic hydrocarbons Chemical class 0.000 claims description 4
- 238000001035 drying Methods 0.000 claims description 4
- 239000003208 petroleum Substances 0.000 claims description 4
- 229910052706 scandium Inorganic materials 0.000 claims description 4
- 239000002904 solvent Substances 0.000 claims description 4
- 150000001338 aliphatic hydrocarbons Chemical class 0.000 claims description 3
- 239000003795 chemical substances by application Substances 0.000 claims description 3
- QWPPOHNGKGFGJK-UHFFFAOYSA-N hypochlorous acid Chemical compound ClO QWPPOHNGKGFGJK-UHFFFAOYSA-N 0.000 claims description 3
- ACVYVLVWPXVTIT-UHFFFAOYSA-N phosphinic acid Chemical class O[PH2]=O ACVYVLVWPXVTIT-UHFFFAOYSA-N 0.000 claims description 3
- 150000003008 phosphonic acid esters Chemical class 0.000 claims description 3
- 150000002910 rare earth metals Chemical class 0.000 claims description 3
- ZMBHCYHQLYEYDV-UHFFFAOYSA-N trioctylphosphine oxide Chemical compound CCCCCCCCP(=O)(CCCCCCCC)CCCCCCCC ZMBHCYHQLYEYDV-UHFFFAOYSA-N 0.000 claims description 3
- HNNQYHFROJDYHQ-UHFFFAOYSA-N 3-(4-ethylcyclohexyl)propanoic acid 3-(3-ethylcyclopentyl)propanoic acid Chemical compound CCC1CCC(CCC(O)=O)C1.CCC1CCC(CCC(O)=O)CC1 HNNQYHFROJDYHQ-UHFFFAOYSA-N 0.000 claims description 2
- 150000001298 alcohols Chemical class 0.000 claims description 2
- 229910000147 aluminium phosphate Inorganic materials 0.000 claims description 2
- 238000002156 mixing Methods 0.000 claims description 2
- 150000003014 phosphoric acid esters Chemical class 0.000 claims description 2
- 238000004090 dissolution Methods 0.000 claims 1
- 229910052692 Dysprosium Inorganic materials 0.000 abstract description 16
- 229910052779 Neodymium Inorganic materials 0.000 abstract description 14
- 229910052791 calcium Inorganic materials 0.000 abstract description 8
- GWEVSGVZZGPLCZ-UHFFFAOYSA-N Titan oxide Chemical compound O=[Ti]=O GWEVSGVZZGPLCZ-UHFFFAOYSA-N 0.000 abstract 2
- 230000002378 acidificating effect Effects 0.000 abstract 1
- 239000012074 organic phase Substances 0.000 description 79
- 239000008346 aqueous phase Substances 0.000 description 40
- 239000012535 impurity Substances 0.000 description 34
- 239000011575 calcium Substances 0.000 description 27
- 239000013078 crystal Substances 0.000 description 14
- 238000001556 precipitation Methods 0.000 description 14
- 235000011121 sodium hydroxide Nutrition 0.000 description 13
- 230000000052 comparative effect Effects 0.000 description 12
- XLYOFNOQVPJJNP-UHFFFAOYSA-M hydroxide Chemical compound [OH-] XLYOFNOQVPJJNP-UHFFFAOYSA-M 0.000 description 11
- 238000003756 stirring Methods 0.000 description 9
- 239000012071 phase Substances 0.000 description 8
- 238000005406 washing Methods 0.000 description 8
- 229910052727 yttrium Inorganic materials 0.000 description 7
- KWYUFKZDYYNOTN-UHFFFAOYSA-M Potassium hydroxide Chemical compound [OH-].[K+] KWYUFKZDYYNOTN-UHFFFAOYSA-M 0.000 description 6
- 150000001875 compounds Chemical class 0.000 description 6
- 230000007423 decrease Effects 0.000 description 6
- QGZKDVFQNNGYKY-UHFFFAOYSA-N Ammonia Chemical compound N QGZKDVFQNNGYKY-UHFFFAOYSA-N 0.000 description 4
- 239000006227 byproduct Substances 0.000 description 4
- WEUCVIBPSSMHJG-UHFFFAOYSA-N calcium titanate Chemical compound [O-2].[O-2].[O-2].[Ca+2].[Ti+4] WEUCVIBPSSMHJG-UHFFFAOYSA-N 0.000 description 4
- 238000010828 elution Methods 0.000 description 4
- 229910052710 silicon Inorganic materials 0.000 description 4
- LSNNMFCWUKXFEE-UHFFFAOYSA-N Sulfurous acid Chemical compound OS(O)=O LSNNMFCWUKXFEE-UHFFFAOYSA-N 0.000 description 3
- WNROFYMDJYEPJX-UHFFFAOYSA-K aluminium hydroxide Chemical compound [OH-].[OH-].[OH-].[Al+3] WNROFYMDJYEPJX-UHFFFAOYSA-K 0.000 description 3
- 150000001768 cations Chemical class 0.000 description 3
- 229910001172 neodymium magnet Inorganic materials 0.000 description 3
- 150000003891 oxalate salts Chemical class 0.000 description 3
- 238000011403 purification operation Methods 0.000 description 3
- 239000011734 sodium Substances 0.000 description 3
- 239000002699 waste material Substances 0.000 description 3
- 150000007513 acids Chemical class 0.000 description 2
- 229910021529 ammonia Inorganic materials 0.000 description 2
- AXCZMVOFGPJBDE-UHFFFAOYSA-L calcium dihydroxide Chemical compound [OH-].[OH-].[Ca+2] AXCZMVOFGPJBDE-UHFFFAOYSA-L 0.000 description 2
- 239000000920 calcium hydroxide Substances 0.000 description 2
- 229910001861 calcium hydroxide Inorganic materials 0.000 description 2
- 230000006866 deterioration Effects 0.000 description 2
- 230000004907 flux Effects 0.000 description 2
- 239000002440 industrial waste Substances 0.000 description 2
- 229910052500 inorganic mineral Inorganic materials 0.000 description 2
- 238000002803 maceration Methods 0.000 description 2
- 239000000696 magnetic material Substances 0.000 description 2
- 235000010755 mineral Nutrition 0.000 description 2
- 239000011707 mineral Substances 0.000 description 2
- 229920006395 saturated elastomer Polymers 0.000 description 2
- 229910052708 sodium Inorganic materials 0.000 description 2
- ZDFBXXSHBTVQMB-UHFFFAOYSA-N 2-ethylhexoxy(2-ethylhexyl)phosphinic acid Chemical compound CCCCC(CC)COP(O)(=O)CC(CC)CCCC ZDFBXXSHBTVQMB-UHFFFAOYSA-N 0.000 description 1
- LJKDOMVGKKPJBH-UHFFFAOYSA-M 2-ethylhexyl hydrogen phosphate Chemical compound CCCCC(CC)COP(O)([O-])=O LJKDOMVGKKPJBH-UHFFFAOYSA-M 0.000 description 1
- 229910018072 Al 2 O 3 Inorganic materials 0.000 description 1
- DGAQECJNVWCQMB-PUAWFVPOSA-M Ilexoside XXIX Chemical compound C[C@@H]1CC[C@@]2(CC[C@@]3(C(=CC[C@H]4[C@]3(CC[C@@H]5[C@@]4(CC[C@@H](C5(C)C)OS(=O)(=O)[O-])C)C)[C@@H]2[C@]1(C)O)C)C(=O)O[C@H]6[C@@H]([C@H]([C@@H]([C@H](O6)CO)O)O)O.[Na+] DGAQECJNVWCQMB-PUAWFVPOSA-M 0.000 description 1
- 229910019142 PO4 Inorganic materials 0.000 description 1
- 229910004298 SiO 2 Inorganic materials 0.000 description 1
- 239000003082 abrasive agent Substances 0.000 description 1
- 239000012670 alkaline solution Substances 0.000 description 1
- DIZPMCHEQGEION-UHFFFAOYSA-H aluminium sulfate (anhydrous) Chemical compound [Al+3].[Al+3].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O DIZPMCHEQGEION-UHFFFAOYSA-H 0.000 description 1
- ANBBXQWFNXMHLD-UHFFFAOYSA-N aluminum;sodium;oxygen(2-) Chemical compound [O-2].[O-2].[Na+].[Al+3] ANBBXQWFNXMHLD-UHFFFAOYSA-N 0.000 description 1
- 150000001450 anions Chemical class 0.000 description 1
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical group [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 description 1
- 230000008901 benefit Effects 0.000 description 1
- QUXFOKCUIZCKGS-UHFFFAOYSA-N bis(2,4,4-trimethylpentyl)phosphinic acid Chemical compound CC(C)(C)CC(C)CP(O)(=O)CC(C)CC(C)(C)C QUXFOKCUIZCKGS-UHFFFAOYSA-N 0.000 description 1
- 239000003054 catalyst Substances 0.000 description 1
- 238000005119 centrifugation Methods 0.000 description 1
- IKNAJTLCCWPIQD-UHFFFAOYSA-K cerium(3+);lanthanum(3+);neodymium(3+);oxygen(2-);phosphate Chemical compound [O-2].[La+3].[Ce+3].[Nd+3].[O-]P([O-])([O-])=O IKNAJTLCCWPIQD-UHFFFAOYSA-K 0.000 description 1
- 230000008859 change Effects 0.000 description 1
- 238000006243 chemical reaction Methods 0.000 description 1
- 239000002734 clay mineral Substances 0.000 description 1
- 239000012043 crude product Substances 0.000 description 1
- 238000010908 decantation Methods 0.000 description 1
- 230000000694 effects Effects 0.000 description 1
- 238000004993 emission spectroscopy Methods 0.000 description 1
- 210000000416 exudates and transudate Anatomy 0.000 description 1
- 239000011521 glass Substances 0.000 description 1
- QOSATHPSBFQAML-UHFFFAOYSA-N hydrogen peroxide;hydrate Chemical compound O.OO QOSATHPSBFQAML-UHFFFAOYSA-N 0.000 description 1
- 238000009616 inductively coupled plasma Methods 0.000 description 1
- 229910052590 monazite Inorganic materials 0.000 description 1
- 230000001590 oxidative effect Effects 0.000 description 1
- TWNQGVIAIRXVLR-UHFFFAOYSA-N oxo(oxoalumanyloxy)alumane Chemical compound O=[Al]O[Al]=O TWNQGVIAIRXVLR-UHFFFAOYSA-N 0.000 description 1
- 229910052760 oxygen Inorganic materials 0.000 description 1
- 239000001301 oxygen Substances 0.000 description 1
- 239000010452 phosphate Substances 0.000 description 1
- UXBZSSBXGPYSIL-UHFFFAOYSA-N phosphoric acid;yttrium(3+) Chemical compound [Y+3].OP(O)(O)=O UXBZSSBXGPYSIL-UHFFFAOYSA-N 0.000 description 1
- 239000003495 polar organic solvent Substances 0.000 description 1
- 230000001376 precipitating effect Effects 0.000 description 1
- 238000002360 preparation method Methods 0.000 description 1
- 238000003825 pressing Methods 0.000 description 1
- 238000002203 pretreatment Methods 0.000 description 1
- 230000009467 reduction Effects 0.000 description 1
- 229910001388 sodium aluminate Inorganic materials 0.000 description 1
- 150000003388 sodium compounds Chemical class 0.000 description 1
- 239000000725 suspension Substances 0.000 description 1
- 230000007704 transition Effects 0.000 description 1
- 229910000164 yttrium(III) phosphate Inorganic materials 0.000 description 1
Images
Classifications
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- 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
- C22B59/00—Obtaining rare earth metals
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B09—DISPOSAL OF SOLID WASTE; RECLAMATION OF CONTAMINATED SOIL
- B09B—DISPOSAL OF SOLID WASTE NOT OTHERWISE PROVIDED FOR
- B09B3/00—Destroying solid waste or transforming solid waste into something useful or harmless
- B09B3/80—Destroying solid waste or transforming solid waste into something useful or harmless involving an extraction step
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
- C22B3/065—Nitric acids or salts thereof
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
- C22B3/10—Hydrochloric acid, other halogenated acids or salts thereof
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
- C22B3/26—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds
- C22B3/38—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds containing phosphorus
- C22B3/384—Pentavalent phosphorus oxyacids, esters thereof
- C22B3/3844—Phosphonic acid, e.g. H2P(O)(OH)2
-
- 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
Definitions
- the present invention leaches rare earth elements, particularly rare earth elements containing Nd and Dy, which are highly useful as materials for Nd—Fe—B based permanent magnets, from rare earth elements containing rare earth elements, and separates and recovers the rare earth elements. More particularly, the present invention relates to a rare earth element recovery method in which rare earth elements are efficiently leached together with Ca from a leaching raw material containing Ca and Ti, and separated and recovered.
- Rare earth elements have been widely used for applications such as phosphors, magnetic materials, abrasives, and catalysts.
- a magnet with a large maximum energy product and residual magnetic flux density can be obtained. It's getting on.
- Patent Document 1 discloses a material for a permanent magnet having an excellent maximum energy product and residual magnetic flux density of an Nd—Fe—B system.
- Patent Document 2 discloses a technique for improving the thermal stability of magnetic properties, which is a disadvantage of the magnet, by replacing part of Nd of the Nd-Fe-B permanent magnet with Dy. ing.
- Such rare earth materials are ores such as monazite, bastonite, xenotime, and ion-adsorbed clay minerals. From these ores, rare earth elements such as mineral acids such as sulfuric acid are used. Although the elements are leached and separated from the obtained leachate, these ore resources are unevenly distributed on the earth, and the abundance ratio of each element in the rare earth elements varies greatly for each ore. In particular, heavy rare earth elements having atomic numbers of 64 to 71 are few in mines that can extract ores that are industrially highly profitable, and Dy, whose demand is particularly high, is concerned about the depletion of resources.
- bauxite which is abundant in resources, is also included in bauxite, which is an ore resource of aluminum, and it is known that rare earth elements are eluted from this bauxite, separated and recovered (for example, , Refer to paragraph 0003 of Patent Document 3), and further, the remainder of the solid residue obtained by collecting aluminum produced as a by-product in the buyer process when the aluminum is produced from this bauxite through the process of the buyer method and the hall elue process (hereinafter referred to as the following) It is also known as “bauxite residue.” When Fe 2 O 3 is the main component, it is red and generally called “red mud.” (Patent Document 4).
- rare earth elements are stable as compounds such as oxides and hydroxides in an aqueous alkaline solution and do not react with sodium hydroxide solution even when heated and pressurized, in the bauxite residue.
- the rare earth element should be concentrated by the amount of the aluminum component eluted by the sodium hydroxide solution in the buyer process described above, and according to the study by the present inventors, compared with the content of the rare earth element in the bauxite.
- the bauxite residue is an industrial waste when producing aluminum from bauxite, and it is stably produced as a by-product when producing aluminum. Since it is easily available, it is expected to be used as a raw material for rare earth elements.
- Patent Document 4 when the above Patent Document 4 is examined in detail, as shown in Examples 1 and 2, 52.0% Fe 2 O 3 and 6.5% TiO 2 in a dry state. , 18.0% loss on ignition, 12.9% Al 2 O 3 , 2.4% SiO 2 , 1.6% Na 2 O, 5.0% CaO, 0.6% P 2 Using a bauxite residue containing O 5 as a raw material, a leaching operation (leaching or digestion) at 10 to 70 ° C.
- Example 1 of Patent Document 4 shows the results of additional testing of Example 1 of Patent Document 4 in which the same leaching operation is repeated three times under the conditions of a temperature of 30 ° C., a pressure of 0.1 MPa, and a time of 15 minutes. Then, the leaching rate of Y remained at 5% by mass or less, and the total leaching rate of Y by the second and third leaching operations was 52% by mass. However, the leaching rates of Nd and Dy were 41% by mass, respectively. It was only 43% by mass, which was still lower than the leaching rate of Y.
- the present inventors have examined the cause of the low leaching rate of rare earth elements, particularly Nd and Dy, in the leaching operation of rare earth elements contained in the bauxite residue, and have reached the following conclusion.
- the bauxite when producing aluminum using bauxite as a raw material, in the buyer process, the bauxite is mixed with an aqueous sodium hydroxide solution, heated and pressurized, and the aluminum component is eluted as aluminate ions, and the obtained aluminum component The eluate containing is cooled, the aluminate ions are precipitated as aluminum hydroxide, and calcined and collected as aluminum oxide.
- CaO is often added to recover the sodium compound produced by the reaction of the components in the bauxite and the aqueous sodium hydroxide solution as sodium hydroxide and to remove impurities such as Si and P in the buyer process.
- the bauxite residue usually contains 4 to 15% by mass of CaO.
- the present inventors also obtained rare earth elements such as Nd and Dy incorporated into the crystal from a leaching raw material containing a compound such as calcium titanate that forms such a perovskite (ABX3) type crystal.
- a compound such as calcium titanate that forms such a perovskite (ABX3) type crystal.
- the perovskite (ABX3) type crystal can be easily dissolved by digestion or maceration, and the rare earth elements not incorporated in the crystal can be dissolved in the crystal. It has been found that the incorporated rare earth elements can be easily leached, and the present invention has been completed.
- an object of the present invention is to provide a rare earth element recovery method in which rare earth elements, particularly rare earth elements including Nd and Dy are efficiently leached from a leaching raw material containing rare earth elements, and separated and recovered.
- an aqueous acid solution is further added and mixed to adjust pH, and the resulting slurry is subjected to a predetermined condition.
- a leaching process in which the leaching process of transferring the rare earth element in the leaching raw material into the acid aqueous solution is carried out, and then the slurry after the leaching process is solid-liquid separated to obtain a leachate containing the rare earth element, and in this leaching process And a separation step of separating and recovering the rare earth element from the obtained leachate, wherein the leaching raw material is dried at 110 ° C.
- S Contains Ca as CaO in a proportion of 4 to 15% by mass and Ti as TiO 2 in a proportion of 2 to 13% by mass, wherein the acid aqueous solution is an acid aqueous solution of hydrochloric acid and / or nitric acid, the preparation PH is 0 ⁇ 7 and the leaching process performed in the leaching step is digestion or maceration performed under heating and pressurizing conditions of a temperature of 160 to 300 ° C. and a pressure of 0.65 to 10 MPa.
- a rare earth element recovery method characterized by leaching a rare earth element in a leaching raw material together with Ca.
- digestion is preferably performed until the elution rate of Ca contained in the leaching raw material reaches 90% by mass or more, whereby Nd and Dy having high utility value including Y are obtained. It is possible to recover a rare earth element containing at a high leaching rate exceeding 70% by mass.
- the term “rare earth element” is used as a collective term for Y of atomic number 39 and La to Lu of atomic numbers 57 to 71. According to the method of the present invention, Sc of atomic number 21 and Ac to Lr of atomic numbers 89 to 103 are also leached, but the present invention denies that these elements are leached and separated and recovered. is not.
- the leaching raw material containing rare earth elements contains rare earth elements such as Y, Nd, and Dy
- Ca is CaO in a proportion of 4 to 15% by mass
- Ti is 2 in terms of TiO 2.
- it is preferably a bauxite residue by-produced in a buyer process of leaching aluminum from bauxite using an aqueous sodium hydroxide solution.
- it is a bauxite residue containing a rare earth element in an amount of 500 to 10,000 ppm as an oxide in the solid component (S) obtained by drying at 110 ° C. for 2 hours.
- Such bauxite residues are by-produced in a buyer process for collecting aluminum from bauxite, particularly in a buyer process in which sodium components are recovered as sodium hydroxide and CaO is added to remove impurities such as Si and P. So it can be easily obtained in large quantities.
- Ca and Ti in the bauxite residue are considered to form crystals of the perovskite (ABX3) type structure.
- the cation at the A site and the anion at the X site Have a similar size, and a cation having a size smaller than the cation at the A site is located at the B site in the cubic lattice composed of the A site and the X site.
- a and B can dissolve various elements as long as the valence and tolerance factor match.
- the tolerance factor t is expressed by the following equation.
- the present invention is a bauxite residue or the like, which is an industrial waste when producing aluminum from bauxite, and Ca is CaO at a rate of 4 to 15% by mass, and Ti is TiO 2 and 2 to 13% by mass. %, It is possible to efficiently and easily leach out rare earth elements, including Nd and Dy, which have high utility value, from the leaching raw materials contained at a ratio of%, and separate and recover them. Besides being able to be used effectively, concerns such as uneven distribution of raw ore of rare earth elements, fluctuation of the abundance ratio of each element in the rare earth elements for each ore, and depletion of resources can be solved.
- FIG. 1 is a flowchart showing impurity element removal and rare earth element concentration of a leachate by a two-stage solvent extraction method according to Example 53 of the present invention.
- the leaching raw material containing a rare earth element is a bauxite residue.
- an aqueous acid solution is added to the bauxite residue and mixed to prepare a slurry.
- the acid aqueous solution used here is preferably an acid aqueous solution containing hydrochloric acid and / or nitric acid which does not form an insoluble compound with Ca in the bauxite residue even when heated to 160 ° C. or higher.
- the slurry to be prepared has a solid / liquid ratio (L / S) of the solid component (S) and the liquid component (L) of preferably 2 or more and 10 or less, more preferably 2 or more and 10 or less,
- the pH value is preferably 0 or more and 2.7 or less, more preferably 0 or more and 2.5 or less.
- an oxidizing agent is added in an amount of 0.1 to 1 equivalent, preferably 0.15 to 0.4 equivalent, with respect to the Fe component in the bauxite residue. It is good.
- Preferred examples of the oxidizing agent added for this purpose include hydrogen peroxide water and aqueous perchloric acid, and more preferably 30% by mass-hydrogen peroxide water and 70% by mass-perchloric acid. It is an aqueous solution. If the amount added is less than 0.1 equivalent, there is a problem that Fe 2+ ions remain in the leachate until the pH is high, and conversely, if it exceeds 1 equivalent, the effect does not change and is wasted. Arise.
- the slurry thus obtained is held under a predetermined condition to carry out a rare earth element leaching treatment.
- the leaching treatment is performed at a temperature of 160 ° C. to 300 ° C., preferably 180 ° C.
- the digestion is performed at a temperature of 250 ° C. or less and a pressure of 0.65 MPa or more and 10 MPa or less, preferably 1 MPa or more and 5 MPa or less, and a holding time of 30 minutes or more and 160 minutes or less, preferably 40 minutes or more and 120 minutes or less.
- the digestion is performed under such a heating and pressurizing condition because Ca and Ti existing in a predetermined ratio in the bauxite residue exist as a compound that forms a crystal having a perovskite structure.
- rare earth elements having a high utility value such as Nd and Dy are incorporated in such a perovskite type structure crystal, so that the rare earth element is leached by dissolving the perovskite type crystal.
- the temperature during the digestion operation is less than 160 ° C.
- the temperature exceeds 250 ° C. the leaching rate of the rare earth element reaches almost saturation, and further heating above 300 ° C. requires the necessary amount of heat. Increase, deterioration of the pressure vessel, and cost increase.
- the holding time during the digestion operation if it is less than 30 minutes, even if the temperature and pressure are set to necessary and appropriate conditions, stable operation becomes difficult due to the short time, Since the leaching rate cannot be stabilized, it becomes difficult to leach 70% by mass or more of the rare earth element contained in the bauxite residue. On the contrary, the leaching rate of the rare earth element is almost saturated after 160 minutes. .
- rare earth elements contained in the bauxite residue are leached together with Ca. It is preferable to perform digestion using the leaching rate of Ca contained in a larger amount than rare earth elements as an index, and it is desirable to perform digestion until the leaching rate of Ca exceeds 90% by mass. By performing digestion until the leaching rate of Ca exceeds 90% by mass, the rare earth elements in the bauxite residue can be reliably leached to exceed 70% by mass.
- the slurry after the leaching treatment is then solid-liquid separated by means such as filtration, centrifugation, and decantation, and a leachate containing rare earth elements together with Ca is recovered.
- the solid residue generated by the solid-liquid separation is preferably washed with washing water, and the leachate adhering to the solid residue is transferred to the washing water and recovered, and the slurry after the leaching treatment is solid-liquid separated first. Together with the leachate obtained in this way, the leachate is treated in the next separation step. If the amount of the washing water used for washing the solid residue is too small, the leachate adhering to the solid residue cannot be sufficiently recovered. On the other hand, if the amount is too large, the load in the next separation step increases.
- the solid-liquid ratio (L / S) between the solid residue (S) and the washing water (L) is usually in the range of 2 to 10.
- the leaching raw material has been described by taking as an example the case of a bauxite residue that is a solid residue after aluminum hydroxide is eluted from bauxite by the Bayer method, but the leaching raw material used in this leaching process As long as it contains rare earth elements and contains 4 to 15% by mass of Ca and 2 to 13% by mass of Ti, it does not have to be a bauxite residue.
- the leachate obtained in the above leaching step is then transferred to a separation step where the rare earth elements are separated and recovered.
- oxalate precipitation, hydroxide precipitation, and solvent extraction are used as separation methods.
- the leachate in which the amount of Fe and Ti eluted is small, the leachate can be directly treated by the oxalate precipitation method or the solvent extraction method, but when the amount of Al or Fe eluted is large, the solvent extraction method or Since the amount of the chemical used in the oxalate precipitation method increases, it is preferable for reducing the cost to reduce the amount of the leachate by pretreatment.
- the pH value of the leachate obtained in the leaching step is usually in the range of 1 to 3
- a pH adjuster is added to the leachate to adjust the pH value to 4 to 6.
- the Fe and Al hydroxides precipitated by this pH adjustment are removed by solid-liquid separation.
- the pH adjuster used for this purpose is not particularly limited, but sodium hydroxide, potassium hydroxide, calcium hydroxide, ammonia and the like are preferably used.
- an oxidizing agent as necessary to oxidize the Fe 2+ ions in the leachate to Fe 3+ ions, so that insoluble Fe (OH) 3 becomes stable, Fe can be easily separated and removed.
- the oxidizing agent for example, hydrogen peroxide, perchloric acid, permanganic acid, hypochlorous acid and the like can be suitably used.
- hydrogen peroxide is used as the oxidizing agent, the concentration of the oxidizing agent only affects the solid-liquid ratio, so an appropriate concentration can be selected from the handling and cost.
- the leaching raw material is bauxite residue, 30% by mass-hydrogen peroxide solution and 70% by mass-perchloric acid aqueous solution are both used in the bauxite residue.
- the amount is preferably 0.1 to 0.5 equivalent with respect to the Fe component.
- the leachate obtained in the above leaching step or the pH of the leachate is adjusted to precipitate Fe and Al as hydroxides.
- a pH adjuster is further added to the liquid obtained by liquid separation to adjust the pH value to 7 or more, and Ca and rare earth elements are precipitated as hydroxides thereof, and these Ca and rare earth element hydroxides. Is solid-liquid separated and recovered as a crude product.
- the pH adjuster is preferably sodium hydroxide, potassium hydroxide, calcium hydroxide, ammonia or the like, and Ca and rare earth elements are precipitated as hydroxides.
- This is solid-liquid separated and recovered as a rare earth element hydroxide, or for the purpose of lowering the concentration of Al, which is an impurity, to the rare earth element hydroxide deposited, at least five times the equivalent of Al. It is also preferable to add a sodium hydroxide solution to dissolve and remove the Al content as aluminate ions.
- the leachate obtained in the above leaching step or the liquid obtained by adjusting the pH of the leachate and precipitating Fe and Al as hydroxides and solid-liquid separation, the solution An insoluble rare earth oxalate is formed by adding 1.3 to 6 equivalents of oxalic acid in the molar amount of all the rare earth elements present therein, and a solid rare earth oxalate compound is obtained as a rare earth oxalate compound by solid-liquid separation. Collect the recovered material.
- a known method may be used as a solvent extraction method, but as an extractant, phosphate ester (DEHPA, EHPA), phosphonate ester (PC88A), phosphinate ester (Cyanex 272, Cyanex 30) with non-polar organic solvents such as hexane, aliphatic hydrocarbons such as benzene, toluene, alcohols such as octanol, and petroleum fractions such as kerosene.
- non-polar organic solvents such as hexane, aliphatic hydrocarbons such as benzene, toluene, alcohols such as octanol, and petroleum fractions such as kerosene.
- the diluted one can be suitably used. It is also preferable to carry out the recovery of the roughly recovered product by the solvent extraction method in two or more stages. According to the recovery of the roughly recovered product by the solvent extraction method over two or more stages, it becomes possible to separate the rare earth element into each element.
- the leaching raw material is a solid residue (bauxite residue) after elution of aluminum hydroxide from bauxite by the Bayer method, and a crude rare earth element compound (crude recovered product) by solvent extraction from the leaching solution obtained in the above leaching step
- the leaching solution is once adjusted to pH 2.5 to 3.5, the deposited precipitate is removed, and the solvent is extracted as it is or after being adjusted again to pH 1.2 to 2.5. It is preferable.
- an emulsion or suspension hereinafter referred to as “emulsion” formed between the organic phase and the aqueous phase during solvent extraction or the like. Occurrence can be prevented. If the emulsion occurs, it can be removed by filtration. If the pH of the aqueous phase at the time of solvent extraction is less than 1.2, the recovery rate of rare earth elements decreases, which is not preferable.
- Such pH adjustment is preferably performed by adding bauxite residue. If the pH is adjusted by adding bauxite residue, the amount of alkaline chemicals used can be suppressed, and the bauxite residue is a by-product in the buyer process when producing aluminum from bauxite. Cost can be reduced. In addition, when the pH is adjusted by the addition of bauxite residue, the rare earth elements contained in the added bauxite residue are eluted in the leachate. Therefore, the acid aqueous solution used for the leaching treatment should be used effectively.
- the rare earth elements leached from the added bauxite residue can be recovered, and at this time, Ca and Ti are precipitated together with Fe, and the concentration of these elements in the leaching solution is lowered. Efficient rare earth element recovery can be performed.
- the extraction rate of the rare earth element can be kept low, and as a result, the concentration of the rare earth element to be separated and recovered can be increased.
- the extraction time is preferably 5 minutes or less, and more preferably 0.5 to 3 minutes.
- the extraction rate of Al can be kept low, and as a result, the concentration of the rare earth element to be separated and recovered can be increased.
- the extraction time exceeds 5 minutes, the extraction rate of Al increases, and as a result, the concentration of the rare earth element to be separated and recovered decreases.
- Sc is separated into the pre-extracted organic phase, but Sc can be recovered as a solid hydroxide from the pre-extracted organic phase by back-extracting an alkaline aqueous solution of pH 7.5 or higher as a back extractant.
- an alkaline aqueous solution of pH 7.5 or higher as a back extractant since Fe and Ti have already been removed, there is no need to adjust pH when extracting rare earth elements using DEHPA as an extractant.
- emulsion may occur between the organic phase and the aqueous phase during solvent extraction. When the emulsion occurs, the precipitate can be removed by filtration.
- the back extraction it is preferable to use a 2N to 8N hydrochloric acid aqueous solution or a 30 to 70% by mass sulfuric acid aqueous solution as a back extractant.
- the back extraction time is preferably 5 minutes or less, and more preferably 0.5 to 3 minutes.
- the back extraction time is 0.5 to 3 minutes, the extraction rate of Al can be kept low, and as a result, the concentration of the rare earth element to be separated and recovered can be increased.
- the back extraction time exceeds 5 minutes, the extraction rate of Al increases, and as a result, the concentration of the rare earth element to be separated and recovered decreases.
- the rare earth element when an aqueous sulfuric acid solution having a concentration of 30 to 70% by mass is used as the back extractant, the rare earth element is precipitated as a solid sulfate, so that the volume can be made very small.
- the back extraction time is preferably 5 minutes or less, and more preferably 0.5 to 3 minutes.
- the extraction rate of Al can be kept low, and as a result, the concentration of the rare earth element to be separated and recovered can be increased.
- the back extraction time exceeds 5 minutes the extraction rate of Al increases, and as a result, the concentration of the rare earth element to be separated and recovered decreases.
- the rare earth element precipitated as solid sulfate can be recovered by solid-liquid separation.
- Al in the organic phase is recovered as aluminum sulfate by performing back extraction for 120 minutes or more using an aqueous sulfuric acid solution having a concentration of 30 to 70% by mass as a back extractant.
- an aqueous sulfuric acid solution having a concentration of 30 to 70% by mass as a back extractant.
- Sc, Ti, Th accumulated in the used extractant is reduced by back extraction using a 2N-8N hydrochloric acid aqueous solution or an alkaline aqueous solution as a back extractant to regenerate. It can be reused as an extractant.
- esters selected from a mixture with tributyl phosphate and / or trioctylphosphine oxide are aliphatic hydrocarbons such as hexane, aromatic hydrocarbons such as benzene and toluene, and petroleum fractions. It is desirable to carry out by a solvent extraction method using an extractant obtained by diluting with a solvent selected from kerosene.
- the separation by such a solvent extraction method is preferably by a countercurrent multistage solvent extraction method.
- the pH value of the leachate is adjusted to 4 to 6, and the Fe and Al hydroxides precipitated by this pH adjustment are solidified.
- a pH adjuster is further added to adjust the pH value to 7 or more, and the precipitated Ca and rare earth element hydroxide are subjected to solid-liquid separation to recover a roughly recovered product.
- Fe and Al are precipitated as hydroxides directly or by adjusting the pH as in the hydroxide precipitation method, and after solid-liquid separation, oxalic acid is added to make rare earth elements as oxalates.
- Precipitated and recovered as a rare earth element oxalate compound then treated with caustic soda to obtain a crude recovery product as a rare earth element hydroxide, or calcined rare earth element oxalate compound as a rare earth element oxide Collect the crude recovery. Since this crude recovery product is dissolved in hydrochloric acid or nitric acid and then solvent extraction is performed using the extractant, there is an advantage that the amount of expensive extractant used in this solvent extraction can be reduced as much as possible. is there.
- Examples 1 to 8 and Comparative Examples 1 to 5 In the solid component (S) obtained by drying at 110 ° C. for 2 hours as the leaching raw material, Fe was 29.8% by mass, Al was 7.9% by mass, and Ca was 5.8% by mass. , 2.1 mass% of Na, 3.5 mass% of Ti, 2.5 mass% of Si, 0.24 of rare earth elements as the sum of Y of atomic number 39 and La to Lu of atomic numbers 57 to 71 Bauxite residues contained in a mass% proportion were used.
- the leachate obtained in the leaching steps of Examples 1 to 8 and Comparative Examples 1 to 5 in this way was analyzed by ICP-AES (inductively coupled plasma emission spectroscopy), respectively.
- Y, Nd, Dy, and Ca in the leachate were analyzed.
- the element contents of Al, Si, Ti, and Fe were measured, and the leaching rate for each element was determined.
- the leachate obtained in the leaching steps of Examples 1 to 8 in each case contained 70% by mass or more of the rare earth element contained in the bauxite residue as the leaching raw material.
- the sulfuric acid aqueous solution is used as the acid aqueous solution
- the solid-liquid ratio (L / S) is 8.6
- the leaching temperature is In the case of Comparative Example 2 at 50 ° C., in the case of Comparative Example 3 using an aqueous phosphoric acid solution as the aqueous acid solution, in the case of Comparative Example 4 having an initial pH value of 3.0, and using an aqueous hypochlorous acid solution as the aqueous acid solution.
- Comparative Example 5 in which the leaching temperature was 100 ° C., 70% by mass or more of the rare earth element contained in the bauxite residue could not be leached.
- Example 9 to 13 and Comparative Examples 6 to 8 Except that the oxidizing agent shown in Table 3 was added to the aqueous acid solution used in the leaching process in the equivalent amount shown in Table 3 with respect to the amount of Fe in the bauxite residue, the same as in Examples 1 to 8 above. Then, leaching of rare earth elements was performed, and the element contents of Y, Nd, Dy, Ca, Al, Si, Ti, and Fe in the obtained leachate were measured, and the leaching rate for each element was determined. These leaching conditions and results are summarized in Table 3.
- Example 14 Using the leachate having the composition shown in Table 4 obtained in Example 4, removal of impurity elements and concentration of rare earth elements were performed by a solvent extraction method.
- this solvent extraction method first, the pH of the leachate is once set to 3.0, the deposited precipitate is removed, adjusted to 1.5, and then DEHPA is diluted with kerosene to a concentration of 0.8M. The exudate and the extractant are stirred and brought into contact with each other at a liquid ratio of 1: 1 for 3 minutes and separated into an extracted organic phase and an aqueous phase after extraction (water phase after extraction). did.
- the extracted organic phase 6N-hydrochloric acid aqueous solution was used as the back extractant, and the extracted organic phase and the back extractant were contacted with stirring at a liquid ratio of 1: 1 for 3 minutes, followed by liquid-liquid separation and back extraction.
- the organic phase after completion (the organic phase after back extraction) and the back extracted water phase were separated, and the rare earth elements in the extracted organic phase were transferred to the back extracted water phase, separated and recovered.
- the organic phase after back extraction is made by using 0.02N-hydrochloric acid aqueous solution as a back extractant, contacting the organic phase after back extraction with the back extractant at a liquid ratio of 1: 1 for 3 minutes, and then separating the liquid and liquid.
- DEHPA can be reused cyclically as an extractant diluted with kerosene to a concentration of 0.8M. Table 5 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method.
- Example 15 to 18 In the same method as in Example 14, the contact time between the leachate and the extractant was 0.5 minutes, 1 minute, 5 minutes, and 10 minutes, and the other conditions were the same as in Example 14, and the rare earth element was back-extracted water. The phase was separated and recovered. Table 5 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method.
- Example 19 to 23 In the same method as in Example 14, the contact time between the extracted organic phase and the back extractant was 0.5 minutes, 1 minute, 5 minutes, 10 minutes, and 15 minutes, and the other conditions were the same as in Example 14. Rare earth elements were transferred to the back-extracted aqueous phase and separated and recovered. Table 5 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method.
- Example 24 Using the leachate having the composition shown in Table 4 obtained in Example 4, removal of impurity elements and concentration of rare earth elements were performed by a solvent extraction method.
- this solvent extraction method first, the pH of the leachate is once adjusted to 1.75, and then an extractant in which DEHPA is diluted to 0.8 M with kerosene is used, and the leachate and the extractant are mixed at a liquid ratio of 1: 1 for 3 minutes. After stirring and contacting, liquid-liquid separation was performed to separate the extracted organic phase and the aqueous phase after extraction. During the solvent extraction, an emulsion was formed between the organic phase and the aqueous phase. The emulsion was separated to the organic phase side during liquid-liquid separation, and then the organic phase was removed by filtration through a filter.
- the extracted organic phase 6N-hydrochloric acid aqueous solution was used as a back extractant, and the extracted organic phase and the back extractant were stirred and contacted at a liquid ratio of 1: 1 for 3 minutes, followed by liquid-liquid separation and back extraction.
- the organic phase and the back-extracted aqueous phase were separated, and the rare earth element was transferred from the extracted organic phase to the back-extracted aqueous phase to be separated and recovered.
- the organic phase after back extraction is made by using 0.02N-hydrochloric acid aqueous solution as a back extractant, contacting the organic phase after back extraction with the back extractant at a liquid ratio of 1: 1 for 3 minutes, and then separating the liquid and liquid.
- DEHPA can be reused cyclically as an extractant diluted with kerosene to a concentration of 0.8M. Table 5 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method.
- Example 25 The other implementation methods and conditions were the same as in Example 14 except that the pH was adjusted by adding the same bauxite residue as used in Example 4 instead of the addition of the aqueous sodium hydroxide solution, and the rare earth element was back-extracted water. The phase was separated and recovered. At this time, the amount of the added bauxite residue was 0.115 kg with respect to 0.1 kg of the bauxite residue which was the leaching raw material.
- Table 5 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method. However, in the calculation of the recovery rate, the recovery rate for 2.15 times the amount of the bauxite residue which was the leaching raw material is shown in consideration of the rare earth element contained in the bauxite residue used for pH adjustment.
- the recovery rate of rare earth elements is higher when the extraction time is shorter, and the recovery rate of rare earth elements is longer when the back extraction time is longer.
- the recovery rate is high, even with Y having the lowest recovery rate, a recovery rate exceeding 75% by mass can be obtained in 1 minute of back extraction time, and impurities such as Al increase as both the extraction time and back extraction time increase. It turns out that the recovery rate of becomes high.
- Example 24 when emulsion was generated between the organic phase and the aqueous phase during solvent extraction, the recovery rate of rare earth elements was compared with Example 14 in which the extraction time and the back extraction time were the same. Can be seen to be slightly lower.
- the recovery rate is recovery from the bauxite residue which was the leaching raw material. Since the recovery rate is lower than that of Example 14 because it is not as high as the rate, it can be seen that Ca and Ti are precipitated together with Fe, and the concentrations of these elements are greatly reduced. Furthermore, the bauxite residue is a by-product in the buyer process when producing aluminum from bauxite, resulting in cost reduction.
- Example 26 Using the leachate having the composition shown in Table 4 obtained in Example 4, removal of impurity elements and concentration of rare earth elements were performed by a solvent extraction method.
- this solvent extraction method first, the pH of the leachate is once adjusted to 3.0, the deposited precipitate is removed, adjusted to 1.0, and then an extractant obtained by diluting DEHPA to 0.8 M with kerosene is added.
- the leachate and the extractant were used and agitated at a liquid ratio of 1: 1 for 3 minutes and then contacted, followed by liquid-liquid separation to separate the extracted organic phase and the extracted aqueous phase.
- the extracted organic phase 6N-hydrochloric acid aqueous solution was used as a back extractant, and the extracted organic phase and the back extractant were stirred and contacted at a liquid ratio of 1: 1 for 3 minutes, followed by liquid-liquid separation and back extraction.
- the organic phase and the back-extracted aqueous phase were separated, and the rare earth element was transferred from the extracted organic phase to the back-extracted aqueous phase to be separated and recovered.
- the organic phase after back extraction is made by using 0.02N-hydrochloric acid aqueous solution as a back extractant, contacting the organic phase after back extraction with the back extractant at a liquid ratio of 1: 1 for 3 minutes, and then separating the liquid and liquid.
- DEHPA can be reused cyclically as an extractant diluted with kerosene to a concentration of 0.8M. Table 6 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method.
- Examples 27 and 28 In the same method as in Example 26, an extractant obtained by diluting DEHPA to 1.2M with kerosene and an extractant obtained by diluting DEHPA to 1.5M with kerosene were used, and other conditions were the same as in Example 26.
- the rare earth elements were transferred to the back-extracted water phase and separated and recovered. Table 6 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method.
- Example 29 and 30 In the same method as in Example 26, the pH of the leachate was once set to 3.0, the deposited precipitate was removed, and then the pH was again adjusted to 1.5 or 2.0. Other conditions were the same as in Example 26. As described above, the rare earth element was transferred to the back-extracted aqueous phase and separated and recovered. Table 6 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method.
- Example 31 Using the leachate having the composition shown in Table 4 obtained in Example 4, removal of impurity elements and concentration of rare earth elements were performed by a solvent extraction method.
- this solvent extraction method first, the pH of the leachate was once set to 3.0, the deposited precipitate was removed, the pH was adjusted to 2.0 again, and PC88A was diluted to 0.8 M with kerosene.
- the extractant Using the extractant, the leachate and the extractant were stirred and brought into contact with each other at a liquid ratio of 1: 1 for 3 minutes, followed by liquid-liquid separation to separate the extracted organic phase and the extracted aqueous phase.
- the extracted organic phase 6N-hydrochloric acid aqueous solution was used as a back extractant, and the extracted organic phase and the back extractant were stirred and contacted at a liquid ratio of 1: 1 for 3 minutes, followed by liquid-liquid separation and back extraction.
- the organic phase and the back-extracted aqueous phase were separated, and the rare earth element was transferred from the extracted organic phase to the back-extracted aqueous phase to be separated and recovered.
- the organic phase after back extraction is made by using 0.02N-hydrochloric acid aqueous solution as a back extractant, contacting the organic phase after back extraction with the back extractant at a liquid ratio of 1: 1 for 3 minutes, and then separating the liquid and liquid. If purified, it can be reused cyclically as an extractant in which PC88A is diluted to a concentration of 0.8M with kerosene. Table 6 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method.
- Examples 32 to 34 In the same method as in Example 31, an extractant obtained by diluting PC88A with kerosene to a concentration of 0.5 to 1.5M was used, and the other conditions were the same as in Example 31, and the rare earth element was transferred to the back-extracted aqueous phase. Separated and recovered. Table 6 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method.
- Examples 35 to 37 In the same method as in Example 31, the pH of the leachate was once set to 3.0, the deposited precipitate was removed, the pH was adjusted again to 1.5 to 3.0, and other conditions were performed. As in Example 31, the rare earth element was transferred to the back-extracted aqueous phase and separated and recovered. Table 6 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method.
- DEHPA has a higher recovery rate of rare earth elements than PC88A, a lower recovery rate of Al, and the extractant is DEHPA.
- PC88A the higher the pH of the leachate, the higher the recovery rate of both rare earth elements and Al, and when DEHPA is used as the extractant, the higher the concentration, the higher the rare earth elements and Al.
- the recovery rate is high, but when PC88A is used as the extractant, the higher the concentration, the higher the recovery rate of rare earth elements, but the recovery rate of Al may have a maximum point near the concentration of 1.2M. I understand that.
- Examples 38 to 43 Using the leachate having the composition shown in Table 4 obtained in Example 4, removal of impurity elements and concentration of rare earth elements were performed by a solvent extraction method including pre-extraction. In this method, first, the pH of the leachate is once set to 3.0, the deposited precipitate is removed, the pH is adjusted to 1.0 or 1.25 again, and then PC88A is made 0.01 to 0 with kerosene. Using a pre-extractant diluted to a concentration of 0.02 M, the leachate and the pre-extractant were brought into contact with stirring at a liquid ratio of 1: 1 for 3 minutes, and then liquid-liquid separation was performed. Separated.
- the extracted organic phase 6N-hydrochloric acid aqueous solution was used as a back extractant, and the extracted organic phase and the back extractant were stirred and contacted at a liquid ratio of 1: 1 for 3 minutes, followed by liquid-liquid separation and back extraction.
- the organic phase and the back-extracted aqueous phase were separated, and the rare earth element was transferred from the extracted organic phase to the back-extracted aqueous phase to be separated and recovered.
- the organic phase after back extraction is prepared by bringing the organic phase after back extraction and the back extractant into contact with each other at a liquid ratio of 10: 1 by stirring for 3 minutes, followed by liquid-liquid separation.
- DEHPA can be reused cyclically as an extractant diluted with kerosene to a concentration of 0.8M. Table 7 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method.
- Example 44 to 52 Using the leachate having the composition shown in Table 4 obtained in Example 4, removal of impurity elements and concentration of rare earth elements were performed by a solvent extraction method.
- this solvent extraction method first, the pH of the leachate was once set to 3.0, the deposited precipitate was removed, the pH was adjusted to 1.0 again, and DEHPA was diluted to 0.8 M with kerosene.
- the extractant Using the extractant, the leachate and the extractant were stirred and brought into contact with each other at a liquid ratio of 1: 1 for 3 minutes, followed by liquid-liquid separation to separate the extracted organic phase and the extracted aqueous phase.
- the extracted organic phase 50% by mass-sulfuric acid aqueous solution was used as the back extractant, and the extracted organic phase and the back extractant were stirred and contacted at a liquid ratio of 1: 1 to 180 minutes. Since an element containing a rare earth element was precipitated as a solid sulfate, the solid sulfate containing the rare earth element was recovered by solid-liquid separation.
- the organic phase after back extraction is made by using 0.02N-hydrochloric acid aqueous solution as a back extractant, contacting the organic phase after back extraction with the back extractant at a liquid ratio of 1: 1 for 3 minutes, and then separating the liquid and liquid.
- DEHPA can be reused cyclically as an extractant diluted with kerosene to a concentration of 0.8M. Table 8 shows the recovery rates of rare earth elements and impurities recovered by this solvent extraction method.
- Example 53 Using the leachate having the composition shown in Table 4 obtained in Example 4, removal of impurity elements and concentration of rare earth elements were performed by a two-stage solvent extraction method shown in FIG.
- a description will be given with reference to FIG.
- the pH of the leachate (1) is adjusted to 2.0, and then an extraction agent obtained by diluting DEHPA to 0.02M with hexane is used in the extraction operation A (Ext. A).
- the leachate (1) and the extractant were agitated and brought into contact with each other at a liquid ratio of 1: 1 for 3 minutes, followed by liquid-liquid separation to separate the extracted organic phase A (2) and the aqueous phase A (3) after extraction.
- Y and Dy are contained in the extracted organic phase A (2), and rare earth elements up to La-Nd are contained in the aqueous phase A (3) after extraction.
- the extracted organic phase C (12) a 0.1N hydrochloric acid aqueous solution was used as a back extractant, and the liquid ratio of the extracted organic phase C (12) and the back extractant was 1 in the back extraction operation C (R-Ext. C). : After stirring for 5 minutes and making contact, liquid-liquid separation is performed and back extraction is performed to obtain organic phase C (15) and back-extracted aqueous phase C (16), and Ca is removed from extracted organic phase C (12). At the same time, the back-extracted aqueous phase C (16) containing Ca was used as the waste liquid (17).
- Example 54 Using the leachate having the composition shown in Table 4 obtained in Example 4, removal of impurity elements and concentration of rare earth elements were carried out by oxalate precipitation. In this oxalate precipitation method, about 1.5 times as much chemical equivalent of oxalic acid as the rare earth ions contained in this liquid was added to the leachate of Example 4 to precipitate only the rare earth elements as oxalate, The rare earth element oxalate was recovered by liquid separation. Table 9 shows the recovery rate of rare earth elements recovered by this oxalate precipitation method and the concentration of impurities.
- Example 55 Using the leachate having the composition shown in Table 4 obtained in Example 4, removal of impurity elements and concentration of rare earth elements were carried out by hydroxide precipitation.
- this hydroxide precipitation method first, the leachate of Example 4 was adjusted to pH 4.5 where the solubility of Al ions and Fe ions was small and the solubility of rare earth element ions was large, and Al and Fe were converted to hydroxides. After removing the precipitated Al and Fe hydroxides by solid-liquid separation, the pH is raised to 11 by adding caustic soda solution to precipitate rare earth ions as hydroxides, and solid-liquid separation Thus, the rare earth element hydroxide was recovered. Table 9 shows the rare earth element recovery rate and impurity concentration by this hydroxide precipitation method.
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Abstract
Description
この浸出液から希土類元素を分離する分離工程では、分離方法として、蓚酸塩析出法、水酸化物析出法、溶媒抽出法が用いられる。
Fe及びTiの溶出量の少ない本発明においては、浸出液を蓚酸塩析出法又は溶媒抽出法により直接処理することも可能であるが、Al又はFeの溶出量が多い場合には、溶媒抽出法又は蓚酸塩析出法で使用する薬剤の使用量が増加するので、浸出液の量を前処理により減少させることがコスト低減のために好ましい。
溶媒抽出法による粗回収物の回収を2段階以上に亘って実施することも好ましい。2段階以上に亘る溶媒抽出法による粗回収物の回収によれば、希土類元素の各元素への分離も可能となる。
逆抽出剤として2N~8Nの塩酸水溶液を用いる場合、逆抽出時間は、5分間以下とすることが好ましく、0.5~3分間とすることがさらに好ましい。逆抽出時間を0.5~3分間とすると、Alの抽出率を低く保つことができ、結果的に分離回収する希土類元素の濃度を高くすることができる。逆抽出時間が5分間を超えるとAlの抽出率が高くなり、結果的に分離回収する希土類元素の濃度が低下する。
使用済みの抽出剤については、2N~8Nの塩酸水溶液又はアルカリ水溶液を逆抽出剤とする逆抽出を行うことにより、前記使用済みの抽出剤中に蓄積したSc、Ti、Thを低減させ、再生抽出剤として再利用することができる。
このような溶媒抽出法による分離は、向流多段溶媒抽出法によることが好適である。
浸出原料として、110℃及び2時間の乾燥条件で乾燥して得られた固体成分(S)中に、Feを29.8質量%、Alを7.9質量%、Caを5.8質量%、Naを2.1質量%、Tiを3.5質量%、Siを2.5質量%、原子番号39のY及び原子番号57~71のLa~Luの合計としての希土類元素を0.24質量%の割合で含有するボーキサイト残渣を使用した。このようなボーキサイト残渣の約0.1kgを圧力容器に装入し、水を添加してスラリーとした後に、表2に示す固液比(L/S)及び初期pH値となるように、塩酸又は硝酸の水溶液を添加し、混合してボーキサイト残渣のスラリーを調製した。
浸出工程で使用した酸水溶液中に表3に示した酸化剤をボーキサイト残渣中のFe量に対して表3に示した当量だけ添加した以外は、上記の実施例1~8の場合と同様にして、希土類元素の浸出を行い、得られた浸出液中のY、Nd、Dy、Ca、Al、Si、Ti、及びFeの元素含有量を測定し、各元素についての浸出率を求めた。これらの浸出条件及び結果を表3にまとめて示す。
実施例4で得られた表4に示す組成の浸出液を用い、溶媒抽出法により不純物元素の除去と希土類元素の濃縮を実施した。この溶媒抽出法では、先ず、浸出液のpHを一旦3.0とし、析出した析出物を除去してから、1.5に調整し、その後に、ケロシンでDEHPAを0.8M濃度に希釈した抽出剤を用い、浸出液と抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて抽出有機相と抽出終了後の水相(抽出後水相)とに分離した。
逆抽出後有機相は、0.02N-塩酸水溶液を逆抽出剤として、逆抽出後有機相と逆抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて精製すれば、ケロシンでDEHPAを0.8M濃度に希釈した抽出剤として循環的に再利用することができる。
この溶媒抽出法で回収された希土類元素と不純物の回収率を表5に示す。
実施例14と同一の方法において、浸出液と抽出剤との接触時間を0.5分、1分、5分及び10分とし、他の条件は実施例14と同一として、希土類元素を逆抽出水相へと移行させて分離し、回収した。
この溶媒抽出法で回収された希希土類元素と不純物の回収率を表5に示す。
実施例14と同一の方法において、抽出有機相と逆抽出剤との接触時間を0.5分、1分、5分、10分及び15分とし、他の条件は実施例14と同一として、希土類元素を逆抽出水相へと移行させて分離し、回収した。
この溶媒抽出法で回収された希土類元素と不純物の回収率を表5に示す。
実施例4で得られた表4に示す組成の浸出液を用い、溶媒抽出法により不純物元素の除去と希土類元素の濃縮を実施した。この溶媒抽出法では、先ず、浸出液のpHを一旦1.75とした後、ケロシンでDEHPAを0.8M濃度に希釈した抽出剤を用い、浸出液と抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて抽出有機相と抽出後水相とに分離した。溶媒抽出時に有機相と水相の中間に乳濁が生成したが、該乳濁は、液液分離時に有機相側に分離し、その後に有機相をフィルターで濾過して除去した。
逆抽出後有機相は、0.02N-塩酸水溶液を逆抽出剤として、逆抽出後有機相と逆抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて精製すれば、ケロシンでDEHPAを0.8M濃度に希釈した抽出剤として循環的に再利用することができる。
この溶媒抽出法で回収された希土類元素と不純物の回収率を表5に示す。
水酸化ナトリウム水溶液の添加に代えて、実施例4に用いたものと同じボーキサイト残渣の添加によりpH調整をした以外、他の実施方法及び条件は実施例14と同一として、希土類元素を逆抽出水相へと移行させて分離し、回収した。このとき、添加したボーキサイト残渣の量は、浸出原料であったボーキサイト残渣0.1kgに対し、0.115kgであった。
この溶媒抽出法で回収された希土類元素と不純物の回収率を表5に示す。ただし回収率の計算にあたっては、pH調整に使用したボーキサイト残渣中に含有されていた希土類元素を考慮し、浸出原料であったボーキサイト残渣の2.15倍の量に対する回収率を示している。
実施例24によれば、溶媒抽出時に有機相と水相の中間に乳濁が生成した場合には、抽出時間及び逆抽出時間が同一である実施例14と比較して、希土類元素の回収率は、わずかに低いことがわかる。
実施例4で得られた表4に示す組成の浸出液を用い、溶媒抽出法により不純物元素の除去と希土類元素の濃縮を実施した。この溶媒抽出法では、先ず、浸出液のpHを一旦3.0とし、析出した析出物を除去してから、1.0に調整した後、ケロシンでDEHPAを0.8M濃度に希釈した抽出剤を用い、浸出液と抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて抽出有機相と抽出後水相とに分離した。
逆抽出後有機相は、0.02N-塩酸水溶液を逆抽出剤として、逆抽出後有機相と逆抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて精製すれば、ケロシンでDEHPAを0.8M濃度に希釈した抽出剤として循環的に再利用することができる。
この溶媒抽出法で回収された希土類元素と不純物の回収率を表6に示す。
実施例26と同一の方法において、ケロシンでDEHPAを1.2M濃度に希釈した抽出剤、及びケロシンでDEHPAを1.5M濃度に希釈した抽出剤を用い、他の条件は実施例26と同一として、希土類元素を逆抽出水相へと移行させて分離し、回収した。
この溶媒抽出法で回収された希土類元素と不純物の回収率を表6に示す。
実施例26と同一の方法において、浸出液のpHを一旦3.0とし、析出した析出物を除去してから、再びpHを1.5又は2.0とし、他の条件は実施例26と同一として、希土類元素を逆抽出水相へと移行させて分離し、回収した。
この溶媒抽出法で回収された希土類元素と不純物の回収率を表6に示す。
実施例4で得られた表4に示す組成の浸出液を用い、溶媒抽出法により不純物元素の除去と希土類元素の濃縮を実施した。この溶媒抽出法では、先ず、浸出液のpHを一旦3.0とし、析出した析出物を除去してから、再びpHを2.0に調整した後、ケロシンでPC88Aを0.8M濃度に希釈した抽出剤を用い、浸出液と抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて抽出有機相と抽出後水相とに分離した。
逆抽出後有機相は、0.02N-塩酸水溶液を逆抽出剤として、逆抽出後有機相と逆抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて精製すれば、ケロシンでPC88Aを0.8M濃度に希釈した抽出剤として循環的に再利用することができる。
この溶媒抽出法で回収された希土類元素と不純物の回収率を表6に示す。
実施例31と同一の方法において、ケロシンでPC88Aを0.5~1.5M濃度に希釈した抽出剤を用い、他の条件は実施例31と同一として、希土類元素を逆抽出水相へと移行させて分離し、回収した。
この溶媒抽出法で回収された希土類元素と不純物の回収率を表6に示す。
実施例31と同一の方法において、浸出液のpHを一旦3.0とし、析出した析出物を除去してから、再びpHの調整を行って1.5~3.0とし、他の条件は実施例31と同一として、希土類元素を逆抽出水相へと移行させて分離し、回収した。
この溶媒抽出法で回収された希土類元素と不純物の回収率を表6に示す。
実施例4で得られた表4に示す組成の浸出液を用い、前抽出を含む溶媒抽出法により不純物元素の除去と希土類元素の濃縮を実施した。この方法では、先ず、浸出液のpHを一旦3.0とし、析出した析出物を除去してから、再びpHを1.0又は1.25に調整した後、ケロシンでPC88Aを0.01~0.02M濃度に希釈した前抽出剤を用い、浸出液と前抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて前抽出有機相と抽出後水相とに分離した。続いて、回収された前抽出有機相について、ケロシンでDEHPAを0.8M濃度に希釈した抽出剤を用い、抽出有機相と抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて抽出有機相と抽出後水相とに分離した。
逆抽出後有機相は、0.02N-塩酸水溶液を逆抽出剤として、逆抽出後有機相と逆抽出剤とを液比10:1で3分間撹拌して接触させた後、液液分離させて精製すれば、ケロシンでDEHPAを0.8M濃度に希釈した抽出剤として循環的に再利用することができる。
この溶媒抽出法で回収された希土類元素と不純物の回収率を表7に示す。
実施例4で得られた表4に示す組成の浸出液を用い、溶媒抽出法により不純物元素の除去と希土類元素の濃縮を実施した。この溶媒抽出法では、先ず、浸出液のpHを一旦3.0とし、析出した析出物を除去してから、再びpHを1.0に調整した後、ケロシンでDEHPAを0.8M濃度に希釈した抽出剤を用い、浸出液と抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて抽出有機相と抽出後水相とに分離した。
逆抽出後有機相は、0.02N-塩酸水溶液を逆抽出剤として、逆抽出後有機相と逆抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて精製すれば、ケロシンでDEHPAを0.8M濃度に希釈した抽出剤として循環的に再利用することができる。
この溶媒抽出法で回収された希土類元素と不純物の回収率を表8に示す。
実施例4で得られた表4に示す組成の浸出液を用い、図1に示す2段階の溶媒抽出法により不純物元素の除去と希土類元素の濃縮を実施した。以下、図1を参照しながら説明する。
この2段階溶媒抽出法では、先ず、浸出液(1)のpHを2.0に調整した後、ヘキサンでDEHPAを0.02M濃度に希釈した抽出剤を用い、抽出操作A(Ext.A)において浸出液(1)と抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて抽出有機相A(2)と抽出後水相A(3)とに分離した。
このとき、YとDyは抽出有機相A(2)に、La-Ndまでの希土類元素は抽出後水相A(3)に、それぞれ含有される。
逆抽出後有機相A(4)については、2N-塩酸水溶液を逆抽出剤として、精製操作(P)において逆抽出後有機相A(4)と逆抽出剤とを液比1:1で3分間撹拌して接触させた後、液液分離させて精製すれば、ヘキサンでDEHPAを0.02M濃度に希釈した抽出剤として循環的に再利用され、また、使用済みの逆抽出剤は廃液(W)として廃棄される。
逆抽出後有機相B(9)は、図示外の上記精製操作(P)と同様にして、ヘキサンでDEHPAを0.02M濃度に希釈した抽出剤として循環的に再利用される。
逆抽出後有機相D(18)は、図示外の上記精製操作(P)と同様にして、ヘキサンでDEHPAを0.8M濃度に希釈した抽出剤として循環的に再利用される。
この2段階溶媒抽出法で回収された希土類元素の回収率と不純物の濃度を表9に示す。
実施例4で得られた表4に示す組成の浸出液を用い、蓚酸塩析出法により不純物元素の除去と希土類元素の濃縮を実施した。この蓚酸塩析出法においては、実施例4の浸出液について、この液中に含まれる希土類元素イオンの約1.5倍の化学等量の蓚酸を加えて希土類元素のみを蓚酸塩として沈殿させ、固液分離して希土類元素蓚酸塩を回収した。
この蓚酸塩析出法で回収された希土類元素の回収率と不純物の濃度を表9に示す。
実施例4で得られた表4に示す組成の浸出液を用い、水酸化物析出法により不純物元素の除去と希土類元素の濃縮を実施した。この水酸化物析出法においては、先ず、実施例4の浸出液について、AlイオンとFeイオンの溶解度が小さく、希土類元素イオンの溶解度が大きなpH4.5に調整して、AlとFeを水酸化物として沈殿させ、固液分離して沈殿したAl及びFeの水酸化物を除去した後、更に苛性ソーダ液を加えてpHを11まで上昇させ、希土類元素イオンを水酸化物として沈殿させ、固液分離して希土類元素水酸化物を回収した。
この水酸化物析出法による希土類元素の回収率と不純物の濃度を表9に示す。
Claims (27)
- 希土類元素を含む浸出原料に水を添加し混合してスラリーを調製した上で、更に酸水溶液を添加し混合してpHを調整し、得られたスラリーを所定の条件下に保持して浸出原料中の希土類元素を酸水溶液中に移行させる浸出処理を行ない、次いで前記浸出処理後のスラリーを固液分離して希土類元素を含む浸出液を得る浸出工程と、この浸出工程で得られた浸出液から希土類元素を分離して回収する分離工程とを有する希土類元素の回収方法であり、
前記浸出原料が、110℃及び2時間の乾燥条件で乾燥して得られた固体成分(S)中に、CaをCaOとして4~15質量%の割合で含むと共にTiをTiO2として2~13質量%の割合で含んでおり、
前記酸水溶液が塩酸及び/又は硝酸を含む酸水溶液であって調整されるpHが0~2.7であり、かつ、
前記浸出工程で行う浸出処理が温度160~300℃及び圧力0.65~10MPaの加熱加圧条件下で行なう温浸であって、この浸出工程では浸出原料中の希土類元素をCaと共に浸出させることを特徴とする希土類元素の回収方法。 - 前記浸出工程では、浸出原料中に含まれるCaの溶出率が90質量%以上に到達するまで温浸を行う請求項1に記載の希土類元素の回収方法。
- 前記浸出原料が、水酸化ナトリウム水溶液を用いてボーキサイトからアルミニウム分を採取するバイヤー工程で副生したボーキサイト残渣である請求項1又は2に記載の希土類元素の回収方法。
- 前記ボーキサイト残渣は、110℃及び2時間の乾燥条件で乾燥して得られた固体成分(S)中に、希土類元素をその酸化物として500~10000ppmの割合で含む請求項3に記載の希土類元素の回収方法。
- 前記ボーキサイト残渣に酸水溶液を添加して得られたスラリーは、固体成分(S)と液体成分(L)との固液比(L/S)が2~10であって、pH値が0~2.7である請求項3又は4に記載の希土類元素の回収方法。
- 前記ボーキサイト残渣に酸水溶液を添加して調製されたスラリー中に、ボーキサイト残渣中のFe成分に対して酸化剤を0.1~1当量の割合で添加する請求項3~5のいずれかに記載の希土類元素の回収方法。
- 前記スラリー中に添加される酸化剤が過酸化水素水又は過塩素酸水溶液である請求項6に記載の希土類元素の回収方法。
- 前記浸出工程で得られた浸出液にpH調整剤を添加してpH4~6に調整し、このpH調整で析出したFe及びAlの水酸化物を固液分離して除去した後、前記分離工程に供する請求項1~7のいずれかに記載の希土類元素の回収方法。
- 浸出液にpH調整剤を添加してpH4~6に調整するpH調整の際に、過酸化水素、過塩素酸、過マンガン酸、及び次亜塩素酸から選ばれた酸化剤を添加し、浸出液中のFe2+イオンを酸化してFe3+イオンにする請求項8に記載の希土類元素の回収方法。
- 前記希土類元素の分離工程では、前記浸出工程で得られた浸出液、又は、該浸出液をpH調整してFe及びAlを水酸化物として沈殿させ、固液分離して得られた液に、pH調整剤を添加してpH7以上に調整し、このpH調整で析出したCa及び希土類元素の水酸化物を固液分離して粗回収物として回収する請求項1~9のいずれかに記載の希土類元素の回収方法。
- 前記希土類元素の分離工程では、前記浸出工程で得られた浸出液、又は、該浸出液をpH調整してFe及びAlを水酸化物として沈殿させ、固液分離して得られた液に存在する希土類元素の化学当量以上の蓚酸を加えて希土類元素を蓚酸塩として析出させ、更に該蓚酸塩を固液分離して前記希土類元素を粗回収物として回収する、請求項1~9のいずれかに記載の希土類元素の回収方法。
- 前記希土類元素の分離工程では、前記浸出工程で得られた浸出液、又は、該浸出液をpH調整してFe及びAlを水酸化物として沈殿させ、固液分離して得られた液に、リン酸エステル類、ホスホン酸エステル類、ホスフィン酸エステル類、チオホスフィン酸エステル類、及びこれらのエステル類とリン酸トリブチル及び/又はトリオクチルホスフィンオキサイドとの混合物から選ばれたエステル類を、ヘキサンなどの脂肪族炭化水素、ベンゼン、トルエンなどの芳香族炭化水素、オクタノールなどのアルコール、及び石油分留物であるケロシンから選ばれた溶媒で希釈して得られた抽出剤を加えて溶媒抽出法により希土類元素を分離して回収する、請求項1~9のいずれかに記載の希土類元素の回収方法。
- 前記溶媒抽出法による分離工程に先駆けて、浸出液のpH調整時に発生した乳濁を予め濾過により除去する、請求項12に記載の希土類元素の回収方法。
- 前記溶媒抽出法による分離工程に先駆けて、浸出液をpH2.5~3.5に調整し、析出した析出物を除去する、請求項12に記載の希土類元素の回収方法。
- 前記溶媒抽出法による分離工程に先駆けて行うpH調整は、ボーキサイト残渣の添加により行われる、請求項14に記載の希土類元素の回収方法。
- 前記溶媒抽出法の抽出剤がDEHPAである、請求項12~15のいずれかに記載の希土類元素の回収方法。
- 前記溶媒抽出法の抽出剤としてのDEHPAの濃度が0.1~1.5Mである、請求項16に記載の希土類元素の回収方法。
- 前記溶媒抽出法の抽出時間が5分間以下である、請求項12~17のいずれかに記載の希土類元素の回収方法。
- 前記溶媒抽出法の抽出時間が0.5~3分間である、請求項18に記載の希土類元素の回収方法。
- 前記DEHPAを抽出剤として用いる溶媒抽出法に先駆けて、PC88A、リン酸トリブチル又はナフテン酸を前抽出剤として用いる浸出液の前抽出を行い、この浸出液からFe、Sc及びTiを分離除去する、請求項16~19のいずれかに記載の希土類元素の回収方法。
- 前記溶媒抽出法は、逆抽出剤が2N~8Nの塩酸水溶液であると共に、逆抽出時間が5分間以下である、請求項12~20のいずれかに記載の希土類元素の回収方法。
- 前記溶媒抽出法の逆抽出時間が0.5~3分間である、請求項21に記載の希土類元素の回収方法。
- 前記溶媒抽出法で用いる逆抽出剤が濃度30~70質量%の硫酸水溶液であり、希土類元素を固体硫酸塩として回収する、請求項12~20のいずれかに記載の希土類元素の回収方法。
- 前記溶媒抽出法の逆抽出時間が5分間以下である、請求項23に記載の希土類元素の回収方法。
- 前記溶媒抽出法において、使用済みの抽出剤に対して、2N~8Nの塩酸水溶液又はアルカリ水溶液を逆抽出剤とする逆抽出を行い、前記使用済みの抽出剤中に蓄積したSc、Ti、Thを低減させ、再生抽出剤として再利用する、請求項12~24のいずれかに記載の希土類元素の回収方法。
- 前記粗回収物から各元素への分離は、粗回収物を酸水溶液に溶解し、次いでリン酸エステル類、ホスホン酸エステル類、ホスフィン酸エステル類、チオホスフィン酸エステル類、及びこれらのエステル類とリン酸トリブチル及び/又はトリオクチルホスフィンオキサイドとの混合物から選ばれたエステル類を、ヘキサンなどの脂肪族炭化水素、ベンゼン、トルエンなどの芳香族炭化水素、及び石油分留物であるケロシンから選ばれた溶媒で希釈して得られた抽出剤を用いる溶媒抽出法で行なう請求項10~25に記載の希土類元素の回収方法。
- 前記粗回収物から各元素への溶媒抽出法による分離が向流多段溶媒抽出法である請求項26に記載の希土類元素の回収方法。
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