JP6437367B2 - Recovery method of rhenium from molybdenum concentrate - Google Patents
Recovery method of rhenium from molybdenum concentrate Download PDFInfo
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- JP6437367B2 JP6437367B2 JP2015072043A JP2015072043A JP6437367B2 JP 6437367 B2 JP6437367 B2 JP 6437367B2 JP 2015072043 A JP2015072043 A JP 2015072043A JP 2015072043 A JP2015072043 A JP 2015072043A JP 6437367 B2 JP6437367 B2 JP 6437367B2
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- ZOKXTWBITQBERF-UHFFFAOYSA-N Molybdenum Chemical compound [Mo] ZOKXTWBITQBERF-UHFFFAOYSA-N 0.000 title claims description 92
- 229910052750 molybdenum Inorganic materials 0.000 title claims description 92
- 239000011733 molybdenum Substances 0.000 title claims description 92
- 239000012141 concentrate Substances 0.000 title claims description 78
- 238000000034 method Methods 0.000 title claims description 44
- 229910052702 rhenium Inorganic materials 0.000 title claims description 23
- WUAPFZMCVAUBPE-UHFFFAOYSA-N rhenium atom Chemical compound [Re] WUAPFZMCVAUBPE-UHFFFAOYSA-N 0.000 title claims description 23
- 238000011084 recovery Methods 0.000 title description 9
- 238000002386 leaching Methods 0.000 claims description 116
- 239000010949 copper Substances 0.000 claims description 68
- 229910052802 copper Inorganic materials 0.000 claims description 62
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims description 57
- 239000011575 calcium Substances 0.000 claims description 40
- XEEYBQQBJWHFJM-UHFFFAOYSA-N iron Substances [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims description 40
- 229910052791 calcium Inorganic materials 0.000 claims description 31
- 239000007788 liquid Substances 0.000 claims description 29
- OYPRJOBELJOOCE-UHFFFAOYSA-N Calcium Chemical compound [Ca] OYPRJOBELJOOCE-UHFFFAOYSA-N 0.000 claims description 27
- 230000003647 oxidation Effects 0.000 claims description 27
- 238000007254 oxidation reaction Methods 0.000 claims description 27
- 230000002378 acidificating effect Effects 0.000 claims description 25
- 229910052742 iron Inorganic materials 0.000 claims description 25
- VEXZGXHMUGYJMC-UHFFFAOYSA-M Chloride anion Chemical compound [Cl-] VEXZGXHMUGYJMC-UHFFFAOYSA-M 0.000 claims description 21
- 239000007864 aqueous solution Substances 0.000 claims description 18
- 239000007789 gas Substances 0.000 claims description 16
- JPVYNHNXODAKFH-UHFFFAOYSA-N Cu2+ Chemical compound [Cu+2] JPVYNHNXODAKFH-UHFFFAOYSA-N 0.000 claims description 15
- 229910001431 copper ion Inorganic materials 0.000 claims description 15
- -1 iron ions Chemical class 0.000 claims description 15
- DYIZHKNUQPHNJY-UHFFFAOYSA-N oxorhenium Chemical compound [Re]=O DYIZHKNUQPHNJY-UHFFFAOYSA-N 0.000 claims description 15
- 229910003449 rhenium oxide Inorganic materials 0.000 claims description 15
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims description 14
- 239000001301 oxygen Substances 0.000 claims description 14
- 229910052760 oxygen Inorganic materials 0.000 claims description 14
- 238000000926 separation method Methods 0.000 claims description 14
- 230000001590 oxidative effect Effects 0.000 claims description 9
- 230000033116 oxidation-reduction process Effects 0.000 claims description 5
- 229910021607 Silver chloride Inorganic materials 0.000 claims description 3
- HKZLPVFGJNLROG-UHFFFAOYSA-M silver monochloride Chemical compound [Cl-].[Ag+] HKZLPVFGJNLROG-UHFFFAOYSA-M 0.000 claims description 3
- 239000000243 solution Substances 0.000 description 19
- ORTQZVOHEJQUHG-UHFFFAOYSA-L copper(II) chloride Chemical compound Cl[Cu]Cl ORTQZVOHEJQUHG-UHFFFAOYSA-L 0.000 description 12
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 11
- 238000006243 chemical reaction Methods 0.000 description 7
- FAPWRFPIFSIZLT-UHFFFAOYSA-M Sodium chloride Chemical compound [Na+].[Cl-] FAPWRFPIFSIZLT-UHFFFAOYSA-M 0.000 description 6
- 239000000203 mixture Substances 0.000 description 6
- 239000000047 product Substances 0.000 description 6
- 239000002994 raw material Substances 0.000 description 6
- 238000000638 solvent extraction Methods 0.000 description 6
- 229910021591 Copper(I) chloride Inorganic materials 0.000 description 5
- 239000002253 acid Substances 0.000 description 5
- OXBLHERUFWYNTN-UHFFFAOYSA-M copper(I) chloride Chemical compound [Cu]Cl OXBLHERUFWYNTN-UHFFFAOYSA-M 0.000 description 5
- 229940045803 cuprous chloride Drugs 0.000 description 5
- 239000000463 material Substances 0.000 description 5
- 229910021578 Iron(III) chloride Inorganic materials 0.000 description 4
- 229960003280 cupric chloride Drugs 0.000 description 4
- 238000005188 flotation Methods 0.000 description 4
- FBAFATDZDUQKNH-UHFFFAOYSA-M iron chloride Chemical compound [Cl-].[Fe] FBAFATDZDUQKNH-UHFFFAOYSA-M 0.000 description 4
- RBTARNINKXHZNM-UHFFFAOYSA-K iron trichloride Chemical compound Cl[Fe](Cl)Cl RBTARNINKXHZNM-UHFFFAOYSA-K 0.000 description 4
- CWQXQMHSOZUFJS-UHFFFAOYSA-N molybdenum disulfide Chemical compound S=[Mo]=S CWQXQMHSOZUFJS-UHFFFAOYSA-N 0.000 description 4
- 150000001340 alkali metals Chemical class 0.000 description 3
- 238000007664 blowing Methods 0.000 description 3
- 238000004090 dissolution Methods 0.000 description 3
- 239000000428 dust Substances 0.000 description 3
- 230000000694 effects Effects 0.000 description 3
- 238000005363 electrowinning Methods 0.000 description 3
- 150000002500 ions Chemical class 0.000 description 3
- 229910052751 metal Inorganic materials 0.000 description 3
- 229910001510 metal chloride Inorganic materials 0.000 description 3
- 229910052961 molybdenite Inorganic materials 0.000 description 3
- 239000011780 sodium chloride Substances 0.000 description 3
- IJGRMHOSHXDMSA-UHFFFAOYSA-N Atomic nitrogen Chemical compound N#N IJGRMHOSHXDMSA-UHFFFAOYSA-N 0.000 description 2
- VTLYFUHAOXGGBS-UHFFFAOYSA-N Fe3+ Chemical compound [Fe+3] VTLYFUHAOXGGBS-UHFFFAOYSA-N 0.000 description 2
- 229910052783 alkali metal Inorganic materials 0.000 description 2
- 229910052784 alkaline earth metal Inorganic materials 0.000 description 2
- 150000001342 alkaline earth metals Chemical class 0.000 description 2
- OMZSGWSJDCOLKM-UHFFFAOYSA-N copper(II) sulfide Chemical compound [S-2].[Cu+2] OMZSGWSJDCOLKM-UHFFFAOYSA-N 0.000 description 2
- 238000007598 dipping method Methods 0.000 description 2
- 238000004993 emission spectroscopy Methods 0.000 description 2
- 238000011156 evaluation Methods 0.000 description 2
- 238000001704 evaporation Methods 0.000 description 2
- 230000008020 evaporation Effects 0.000 description 2
- 238000000605 extraction Methods 0.000 description 2
- 238000005342 ion exchange Methods 0.000 description 2
- 239000002184 metal Substances 0.000 description 2
- JKQOBWVOAYFWKG-UHFFFAOYSA-N molybdenum trioxide Chemical compound O=[Mo](=O)=O JKQOBWVOAYFWKG-UHFFFAOYSA-N 0.000 description 2
- 238000007781 pre-processing Methods 0.000 description 2
- 238000002203 pretreatment Methods 0.000 description 2
- 230000001737 promoting effect Effects 0.000 description 2
- 238000003756 stirring Methods 0.000 description 2
- 229910052717 sulfur Inorganic materials 0.000 description 2
- 239000002699 waste material Substances 0.000 description 2
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 2
- QGZKDVFQNNGYKY-UHFFFAOYSA-O Ammonium Chemical compound [NH4+] QGZKDVFQNNGYKY-UHFFFAOYSA-O 0.000 description 1
- KZBUYRJDOAKODT-UHFFFAOYSA-N Chlorine Chemical compound ClCl KZBUYRJDOAKODT-UHFFFAOYSA-N 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
- WHXSMMKQMYFTQS-UHFFFAOYSA-N Lithium Chemical compound [Li] WHXSMMKQMYFTQS-UHFFFAOYSA-N 0.000 description 1
- FYYHWMGAXLPEAU-UHFFFAOYSA-N Magnesium Chemical compound [Mg] FYYHWMGAXLPEAU-UHFFFAOYSA-N 0.000 description 1
- ZLMJMSJWJFRBEC-UHFFFAOYSA-N Potassium Chemical compound [K] ZLMJMSJWJFRBEC-UHFFFAOYSA-N 0.000 description 1
- 229910000831 Steel Inorganic materials 0.000 description 1
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 1
- MYOSDXHQLHUZPL-UHFFFAOYSA-N [Pb].O.[Fe] Chemical compound [Pb].O.[Fe] MYOSDXHQLHUZPL-UHFFFAOYSA-N 0.000 description 1
- RQFRTWTXFAXGQQ-UHFFFAOYSA-N [Pb].[Mo] Chemical compound [Pb].[Mo] RQFRTWTXFAXGQQ-UHFFFAOYSA-N 0.000 description 1
- 238000005273 aeration Methods 0.000 description 1
- 239000003570 air Substances 0.000 description 1
- 229910001514 alkali metal chloride Inorganic materials 0.000 description 1
- 229910001617 alkaline earth metal chloride Inorganic materials 0.000 description 1
- 229910052788 barium Inorganic materials 0.000 description 1
- DSAJWYNOEDNPEQ-UHFFFAOYSA-N barium atom Chemical compound [Ba] DSAJWYNOEDNPEQ-UHFFFAOYSA-N 0.000 description 1
- 239000010953 base metal Substances 0.000 description 1
- 229910052790 beryllium Inorganic materials 0.000 description 1
- ATBAMAFKBVZNFJ-UHFFFAOYSA-N beryllium atom Chemical compound [Be] ATBAMAFKBVZNFJ-UHFFFAOYSA-N 0.000 description 1
- 229910052792 caesium Inorganic materials 0.000 description 1
- TVFDJXOCXUVLDH-UHFFFAOYSA-N caesium atom Chemical compound [Cs] TVFDJXOCXUVLDH-UHFFFAOYSA-N 0.000 description 1
- 210000004027 cell Anatomy 0.000 description 1
- 229910052947 chalcocite Inorganic materials 0.000 description 1
- DVRDHUBQLOKMHZ-UHFFFAOYSA-N chalcopyrite Chemical compound [S-2].[S-2].[Fe+2].[Cu+2] DVRDHUBQLOKMHZ-UHFFFAOYSA-N 0.000 description 1
- 229910052951 chalcopyrite Inorganic materials 0.000 description 1
- 239000012295 chemical reaction liquid Substances 0.000 description 1
- 150000001805 chlorine compounds Chemical class 0.000 description 1
- 238000004140 cleaning Methods 0.000 description 1
- 230000000052 comparative effect Effects 0.000 description 1
- 238000007796 conventional method Methods 0.000 description 1
- 230000007423 decrease Effects 0.000 description 1
- 210000001787 dendrite Anatomy 0.000 description 1
- 238000010828 elution Methods 0.000 description 1
- 230000001747 exhibiting effect Effects 0.000 description 1
- 229960002089 ferrous chloride Drugs 0.000 description 1
- 229910052730 francium Inorganic materials 0.000 description 1
- KLMCZVJOEAUDNE-UHFFFAOYSA-N francium atom Chemical compound [Fr] KLMCZVJOEAUDNE-UHFFFAOYSA-N 0.000 description 1
- 239000011521 glass Substances 0.000 description 1
- 238000010438 heat treatment Methods 0.000 description 1
- IXCSERBJSXMMFS-UHFFFAOYSA-N hydrogen chloride Substances Cl.Cl IXCSERBJSXMMFS-UHFFFAOYSA-N 0.000 description 1
- 229910000041 hydrogen chloride Inorganic materials 0.000 description 1
- 239000012535 impurity Substances 0.000 description 1
- 239000011261 inert gas Substances 0.000 description 1
- 229910052500 inorganic mineral Inorganic materials 0.000 description 1
- NMCUIPGRVMDVDB-UHFFFAOYSA-L iron dichloride Chemical compound Cl[Fe]Cl NMCUIPGRVMDVDB-UHFFFAOYSA-L 0.000 description 1
- 229910052744 lithium Inorganic materials 0.000 description 1
- 229910052749 magnesium Inorganic materials 0.000 description 1
- 239000011777 magnesium Substances 0.000 description 1
- 239000011707 mineral Substances 0.000 description 1
- 239000011259 mixed solution Substances 0.000 description 1
- 229910052757 nitrogen Inorganic materials 0.000 description 1
- 229910052756 noble gas Inorganic materials 0.000 description 1
- 239000007800 oxidant agent Substances 0.000 description 1
- 239000002245 particle Substances 0.000 description 1
- 229910052700 potassium Inorganic materials 0.000 description 1
- 239000011591 potassium Substances 0.000 description 1
- 239000000843 powder Substances 0.000 description 1
- 238000001556 precipitation Methods 0.000 description 1
- 239000010909 process residue Substances 0.000 description 1
- 238000012545 processing Methods 0.000 description 1
- 229910052705 radium Inorganic materials 0.000 description 1
- HCWPIIXVSYCSAN-UHFFFAOYSA-N radium atom Chemical compound [Ra] HCWPIIXVSYCSAN-UHFFFAOYSA-N 0.000 description 1
- 238000011160 research Methods 0.000 description 1
- 239000011347 resin Substances 0.000 description 1
- 229920005989 resin Polymers 0.000 description 1
- 229910052701 rubidium Inorganic materials 0.000 description 1
- IGLNJRXAVVLDKE-UHFFFAOYSA-N rubidium atom Chemical compound [Rb] IGLNJRXAVVLDKE-UHFFFAOYSA-N 0.000 description 1
- 150000003839 salts Chemical class 0.000 description 1
- 229910052708 sodium Inorganic materials 0.000 description 1
- 239000011734 sodium Substances 0.000 description 1
- 238000004611 spectroscopical analysis Methods 0.000 description 1
- 238000005507 spraying Methods 0.000 description 1
- 239000010959 steel Substances 0.000 description 1
- 229910052712 strontium Inorganic materials 0.000 description 1
- CIOAGBVUUVVLOB-UHFFFAOYSA-N strontium atom Chemical compound [Sr] CIOAGBVUUVVLOB-UHFFFAOYSA-N 0.000 description 1
- 238000006467 substitution reaction Methods 0.000 description 1
- 150000004763 sulfides Chemical class 0.000 description 1
- 239000011593 sulfur Substances 0.000 description 1
- 238000012360 testing method Methods 0.000 description 1
- 239000002562 thickening agent Substances 0.000 description 1
- 238000005406 washing Methods 0.000 description 1
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- 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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- Manufacture And Refinement Of Metals (AREA)
Description
本発明はモリブデン精鉱からのレニウムの回収方法に関する。 The present invention relates to a method for recovering rhenium from molybdenum concentrate.
レニウムは、銅鉱に付随して産出される輝水鉛鉱（ｍｏｌｙｂｄｅｎｉｔｅ，ＭｏＳ2）等のモリブデン鉱に含まれており、モリブデン鉱からのモリブデン回収の際に、レニウムも揮発させて回収することが従来から行われている。特に、モリブデン鉱中に含まれるレニウムの酸化物（特にＲｅ2Ｏ7）は揮発性であるため、浮遊選鉱法により銅とモリブデンを選別して得られたモリブデン精鉱を酸化焙焼処理し、酸化焙焼処理によって揮発した酸化レニウム（Ｒｅ2Ｏ7）を回収する方法が知られている（例えば、非特許文献１参照）。 Rhenium is contained in molybdenum ores such as molybdenite (MoS 2 ) produced in association with copper ore, and when recovering molybdenum from molybdenum ore, rhenium is also volatilized and recovered. It is made from. In particular, the oxide of rhenium (especially Re 2 O 7 ) contained in molybdenum ore is volatile, so the molybdenum concentrate obtained by sorting copper and molybdenum by the flotation process is oxidized and roasted, A method of recovering rhenium oxide (Re 2 O 7 ) volatilized by oxidative roasting is known (see, for example, Non-Patent Document 1).
モリブデン鉱は、銅の硫化物とともに産出されることが多いために、浮遊選鉱法により銅とモリブデンを選別した後、通常回収したモリブデン精鉱中には数％の銅の硫化物が混在している。銅は鉄鋼製品の性状を低下させるため、塩化鉄法によってモリブデン精鉱から予め銅を除去することが一般的に行われている。 Molybdenum ore is often produced together with copper sulfide, so after separating copper and molybdenum by the flotation method, usually several percent of copper sulfide is mixed in the recovered molybdenum concentrate. Yes. Since copper deteriorates the properties of steel products, it is generally performed in advance to remove copper from molybdenum concentrate by the iron chloride method.
例えば、米国特許第３６７４４２４号明細書（特許文献１）には、アルカリ金属又はアルカリ土類金属の塩化物を含有する水溶液を浸出液として使用し、モリブデン精鉱から銅を除去して純度を高める方法が記載されている。 For example, US Pat. No. 3,674,424 (Patent Document 1) uses an aqueous solution containing an alkali metal or alkaline earth metal chloride as a leachate and removes copper from molybdenum concentrate to increase the purity. Is described.
特開昭５８−１８９３４３号公報（特許文献２）には、モリブデン鉱選鉱物内に存在する不純物として含まれる銅を、塩化第二鉄を用いて浸出する方法が記載されている。 Japanese Patent Laid-Open No. 58-189343 (Patent Document 2) describes a method of leaching copper contained as an impurity present in a molybdenum ore mineral using ferric chloride.
しかしながら、特許文献１及び２に記載されるような塩化鉄法を用いた場合には、アルカリ金属又はアルカリ土類金属の塩化物を含有する塩素ガスや高圧ガスを使用した高温処理を行うために処理に危険を伴い、コストや安全面での課題が残る。更に、特許文献１及び２には、モリブデン精鉱から銅を除去する具体的方法についての記載があるのみで、その後の焙焼処理により揮発して得られる酸化レニウムの回収に関する知見は一切されていない。 However, in the case of using the iron chloride method as described in Patent Documents 1 and 2, in order to perform high temperature treatment using chlorine gas or high pressure gas containing chloride of alkali metal or alkaline earth metal There is a risk in processing, and there are still problems in cost and safety. Furthermore, Patent Documents 1 and 2 only describe a specific method for removing copper from molybdenum concentrate, and no knowledge about the recovery of rhenium oxide obtained by volatilization by subsequent roasting treatment has been made. Absent.
一方、非特許文献１には、酸化焙焼後のレニウムの回収方法の概要については記載されているが、レニウムの揮発率を向上させるための具体的方策については、記載も示唆もされていない。 On the other hand, Non-Patent Document 1 describes an outline of a method for recovering rhenium after oxidative roasting, but does not describe or suggest a specific measure for improving the volatilization rate of rhenium. .
上記課題を鑑み、本発明は、従来に比べてより生産性が高く安全な方法で、酸化レニウムの揮発率を向上させることが可能なモリブデン精鉱からのレニウムの回収方法を提供する。 In view of the above problems, the present invention provides a method for recovering rhenium from molybdenum concentrate that can improve the volatility of rhenium oxide by a safer and more productive method than conventional methods.
本発明者は鋭意検討を重ねた結果、酸性塩化浴、即ち塩化物イオンを含む酸性水溶液を用いてモリブデン精鉱中の銅とカルシウムとを除去する前処理を施し、前処理を施したモリブデン精鉱に対して、酸化焙焼処理をすることで、より生産性が高く安全な方法を用いて、酸化レニウムの揮発率を向上させることが可能となることを見いだした。 As a result of intensive studies, the present inventor performed a pretreatment for removing copper and calcium in molybdenum concentrate using an acidic chloride bath, that is, an acidic aqueous solution containing chloride ions, and the molybdenum concentrate subjected to the pretreatment. It has been found that by subjecting the ore to oxidative roasting, it is possible to improve the volatility of rhenium oxide using a safer and more productive method.
以上の知見を基礎として完成した本発明は一側面において、塩化物イオンを含む酸性水溶液を用いてモリブデン精鉱に含まれる銅及びカルシウムを浸出する前処理工程と、前処理工程で得られたモリブデン精鉱処理物を酸化焙焼する酸化焙焼工程と、酸化焙焼工程において揮発した酸化レニウムを回収する工程とを含むモリブデン精鉱からのレニウムの回収方法が提供される。 The present invention completed on the basis of the above knowledge is, in one aspect, a pretreatment step of leaching copper and calcium contained in molybdenum concentrate using an acidic aqueous solution containing chloride ions, and molybdenum obtained in the pretreatment step. There is provided a method for recovering rhenium from molybdenum concentrate, which includes an oxidation roasting step for oxidizing and roasting the concentrate concentrate and a step of recovering rhenium oxide volatilized in the oxidation roasting step.
本発明に係るモリブデン精鉱からのレニウムの回収方法は一実施態様において、塩化物イオンを含む酸性水溶液が、銅イオン又は鉄イオンの少なくともいずれかを含む酸性水溶液であり、前処理工程が、酸性水溶液に酸化剤を供給しながらモリブデン精鉱中の銅及びカルシウムを浸出し、次いで固液分離により浸出残渣と浸出後液を得ることを含む。 In one embodiment of the method for recovering rhenium from molybdenum concentrate according to the present invention, the acidic aqueous solution containing chloride ions is an acidic aqueous solution containing at least one of copper ions or iron ions, and the pretreatment step is acidic. It includes leaching copper and calcium in molybdenum concentrate while supplying an oxidizing agent to an aqueous solution, and then obtaining a leaching residue and a liquid after leaching by solid-liquid separation.
本発明に係るモリブデン精鉱からのレニウムの回収方法は別の一実施態様において、前処理工程が、塩化物イオンを１００〜２００ｇ／Ｌ、銅イオンを１〜３０ｇ／Ｌ、鉄イオンを１〜１０ｇ／Ｌ含有する５０〜１００℃の酸性水溶液に酸素含有気体を供給しながらモリブデン精鉱を接触させてモリブデン精鉱中の銅及びカルシウムを浸出する工程と、次いで固液分離により浸出後液と浸出残渣を得ることを含む。 In another embodiment of the method for recovering rhenium from the molybdenum concentrate according to the present invention, the pretreatment step comprises chloride ions of 100 to 200 g / L, copper ions of 1 to 30 g / L, and iron ions of 1 to 1. A step of leaching copper and calcium in the molybdenum concentrate by contacting the molybdenum concentrate while supplying an oxygen-containing gas to an acidic aqueous solution containing 10 g / L at 50 to 100 ° C., and then a liquid after leaching by solid-liquid separation Including obtaining a leaching residue.
本発明に係るモリブデン精鉱からのレニウムの回収方法は更に別の一実施態様において、前処理工程が、塩化物イオンを１００〜２００ｇ／Ｌ、銅イオンを１〜３０ｇ／Ｌ、鉄イオンを１〜２０ｇ／Ｌ含有する５０〜１００℃の酸性水溶液をモリブデン精鉱に接触させて銅及びカルシウムを浸出する工程と、次いで固液分離により浸出後液と浸出残渣を得る工程を含み、銅及びカルシウムの浸出終点における酸性水溶液の酸化還元電位（ｖｓ Ａｇ／ＡｇＣｌ）を４００〜４８０ｍＶとすることを含む。 In yet another embodiment of the method for recovering rhenium from molybdenum concentrate according to the present invention, the pretreatment step comprises 100 to 200 g / L of chloride ions, 1 to 30 g / L of copper ions, and 1 of iron ions. A step of leaching copper and calcium by contacting an acidic aqueous solution containing ˜20 g / L at 50 to 100 ° C. with molybdenum concentrate, and then obtaining a post-leaching solution and a leaching residue by solid-liquid separation; The oxidation-reduction potential (vs Ag / AgCl) of the acidic aqueous solution at the end point of leaching of 400 to 480 mV.
本発明に係るモリブデン精鉱からのレニウムの回収方法は更に別の一実施態様において、モリブデン精鉱が、銅を０．５〜１０質量％、カルシウムを０．１５〜５．０質量％、鉄を０．３〜１０質量％、モリブデンを２０〜６０質量％含有する。 In yet another embodiment of the method for recovering rhenium from molybdenum concentrate according to the present invention, the molybdenum concentrate contains 0.5 to 10% by mass of copper, 0.15 to 5.0% by mass of calcium, and iron. 0.3 to 10% by mass and 20 to 60% by mass of molybdenum.
本発明によれば、従来に比べてより生産性が高く安全な方法で、モリブデン精鉱の酸化焙焼処理で得られる酸化レニウムの揮発率を向上させることが可能なモリブデン精鉱からのレニウムの回収方法が得られる。 According to the present invention, rhenium from a molybdenum concentrate that can improve the volatility of rhenium oxide obtained by oxidation roasting treatment of molybdenum concentrate in a more productive and safer method than in the past. A recovery method is obtained.
以下、本発明の具体的な実施形態について詳細に説明するが、本発明は以下の実施形態に制限されるものではない。 Hereinafter, specific embodiments of the present invention will be described in detail, but the present invention is not limited to the following embodiments.
本発明の実施の形態に係るモリブデン精鉱からのレニウムの回収方法は、（１）モリブデン精鉱に含まれる銅及びカルシウムを浸出する前処理工程と、（２）前処理工程で得られたモリブデン精鉱処理物を酸化焙焼する酸化焙焼工程と、（３）酸化焙焼工程において揮発した酸化レニウムを回収する工程とを含む。 The method for recovering rhenium from molybdenum concentrate according to the embodiment of the present invention includes (1) a pretreatment step of leaching copper and calcium contained in molybdenum concentrate, and (2) molybdenum obtained in the pretreatment step. An oxidation roasting step of oxidizing and roasting the concentrate processed product, and (3) a step of recovering rhenium oxide volatilized in the oxidation roasting step.
本実施形態において処理対象とするモリブデン精鉱としては、レニウムを含むモリブデン精鉱であれば特に制限されず、例えば輝水鉛鉱、モリブデン鉛鉱、パウエライト、及び鉄水鉛鉱から選択される一種以上を含有する鉱石、とりわけ輝水鉛鉱を含有する鉱石を浮遊選鉱した後のモリブデン精鉱が好適に用いられる。 In the present embodiment, the molybdenum concentrate to be treated is not particularly limited as long as it is a molybdenum concentrate containing rhenium. For example, at least one selected from molybdenite, molybdenum lead, pawerite, and iron-water lead ore is used. Molybdenum concentrates obtained by flotation of ores containing, especially ores containing molybdenite, are preferably used.
モリブデン精鉱中の銅は硫化物の形態、例えば輝銅鉱及び／又は黄銅鉱の形態で存在してもよい。モリブデン精鉱中の成分組成は浮選条件等によっても異なるが、一実施形態においてはモリブデン精鉱がＣｕを０．５〜１０質量％含有し、典型的な一実施形態においてＣｕを１〜１０質量％含有し、より典型的な一実施形態においてＣｕを２〜５質量％含有する。モリブデン精鉱は一実施形態においてＭｏを２０質量％以上含有し、典型的にはＭｏを２０〜６０質量％含有し、より典型的にはＭｏを３０〜６０質量％含有し、更に典型的な一実施形態においてＭｏを４０〜５０質量％含有する。 The copper in the molybdenum concentrate may be present in the form of sulfides, for example in the form of chalcocite and / or chalcopyrite. Although the component composition in the molybdenum concentrate varies depending on the flotation conditions and the like, in one embodiment, the molybdenum concentrate contains 0.5 to 10% by mass of Cu, and in one typical embodiment, 1 to 10 Cu. In a more typical embodiment, 2 to 5% by mass of Cu is contained. The molybdenum concentrate in one embodiment contains 20% by mass or more of Mo, typically contains 20-60% by mass of Mo, more typically contains 30-60% by mass of Mo, and is more typical. In one Embodiment, Mo is contained 40-50 mass%.
更に、モリブデン精鉱は一実施形態においてＣａを０．１５〜５．０質量％含有し、より典型的にはＣａを０．３〜２．０質量％含有する。更に、モリブデン精鉱は一実施形態においてＦｅを０．３〜１０質量％含有する。モリブデン精鉱におけるＲｅ含有量は特に限定されないが、典型的には２００〜７００質量ｐｐｍ含有する。 Furthermore, molybdenum concentrate contains 0.15-5.0 mass% of Ca in one Embodiment, More typically, it contains 0.3-2.0 mass% of Ca. Furthermore, the molybdenum concentrate contains 0.3 to 10% by mass of Fe in one embodiment. The Re content in the molybdenum concentrate is not particularly limited, but typically contains 200 to 700 ppm by mass.
（１）前処理工程
本実施形態に係る前処理工程ではまず、塩化物イオンを含む酸性水溶液（以下、「浸出液」ともいう。）を用いて、モリブデン精鉱中に含まれる銅及びカルシウムを浸出させる浸出工程と、浸出工程で得られた浸出反応液を固液分離する工程を含む。
(1) Pretreatment step In the pretreatment step according to this embodiment, first, copper and calcium contained in molybdenum concentrate are leached using an acidic aqueous solution containing chloride ions (hereinafter also referred to as “leaching solution”). And a step of solid-liquid separation of the leaching reaction liquid obtained in the leaching step.
酸性水溶液への塩化物イオンの供給源としては特に制限はなく、例えば塩化水素、塩酸及び塩化金属等が挙げられる。経済性や安全性を考慮すれば塩化金属の形態で供給するのが好ましい。塩化金属としては、例えば塩化銅（塩化第一銅、塩化第二銅）、アルカリ金属（リチウム、ナトリウム、カリウム、ルビジウム、セシウム、フランシウム）の塩化物、アルカリ土類金属（ベリリウム、マグネシウム、カルシウム、ストロンチウム、バリウム、ラジウム）の塩化物が挙げられ、経済性や入手容易性の観点から、塩化ナトリウムが好ましい。また、銅イオンの供給源としても利用できることから、塩化銅を利用することも好ましい。 There is no restriction | limiting in particular as a supply source of the chloride ion to acidic aqueous solution, For example, hydrogen chloride, hydrochloric acid, a metal chloride, etc. are mentioned. In consideration of economy and safety, it is preferable to supply in the form of metal chloride. Examples of the metal chloride include copper chloride (cuprous chloride, cupric chloride), alkali metal (lithium, sodium, potassium, rubidium, cesium, francium) chloride, alkaline earth metal (beryllium, magnesium, calcium, (Strontium, barium, radium) chlorides are mentioned, and sodium chloride is preferred from the viewpoint of economy and availability. Moreover, since it can utilize also as a supply source of copper ion, it is also preferable to utilize copper chloride.
酸性水溶液は、塩化物イオンに加えて更に銅イオン又は鉄イオンの少なくともいずれかを含むことが好ましい。銅イオン及び鉄イオンは、塩の形態で供給するのが通常であり、例えばハロゲン化塩の形態で供給することができる。塩化物イオンの供給源としても利用できる観点から銅イオンは塩化銅、鉄イオンは塩化鉄として供給されるのが好ましい。塩化銅及び塩化鉄としては酸化力の観点から塩化第二銅（ＣｕＣｌ2）及び塩化第二鉄（ＦｅＣｌ3）を使用するのがそれぞれ望ましいが、塩化第一銅（ＣｕＣｌ）及び塩化第一鉄（ＦｅＣｌ2）を使用しても浸出液に酸素含有気体を供給することで、塩化第二銅（ＣｕＣｌ2）及び塩化第二鉄（ＦｅＣｌ3）にそれぞれ酸化されるため、大差はない。 The acidic aqueous solution preferably further contains at least one of copper ions and iron ions in addition to chloride ions. Copper ions and iron ions are usually supplied in the form of a salt, and can be supplied, for example, in the form of a halide salt. From the viewpoint that it can also be used as a supply source of chloride ions, copper ions are preferably supplied as copper chloride and iron ions are preferably supplied as iron chloride. As copper chloride and iron chloride, it is preferable to use cupric chloride (CuCl 2 ) and ferric chloride (FeCl 3 ) from the viewpoint of oxidizing power, respectively, but cuprous chloride (CuCl) and ferrous chloride are preferable. Even if (FeCl 2 ) is used, supplying an oxygen-containing gas to the leachate oxidizes to cupric chloride (CuCl 2 ) and ferric chloride (FeCl 3 ), respectively, so there is no significant difference.
モリブデン精鉱からその後の焙焼処理において効率良く酸化レニウムを揮発させて回収するためには、浸出工程で使用する浸出液は酸性とすべきである。特に、塩化物イオンの供給源としても利用できることから、塩酸酸性とするのが好ましい。浸出液のｐＨは、浸出した銅及びカルシウムの溶解度を確保する理由から、ガラス電極によって測定されるｐＨを０〜３程度とするのが好ましく、０．２〜２．５程度とするのがより好ましい。本実施形態においては、前処理工程に使用される浸出液として、塩酸、塩化第二銅、塩化第二鉄、及び塩化ナトリウムの混合液を使用することができる。 In order to efficiently volatilize and recover rhenium oxide in the subsequent roasting process from molybdenum concentrate, the leachate used in the leaching process should be acidic. In particular, since it can be used as a supply source of chloride ions, it is preferably made acidic with hydrochloric acid. The pH of the leaching solution is preferably about 0 to 3 and more preferably about 0.2 to 2.5 as measured by the glass electrode for ensuring the solubility of the leached copper and calcium. . In the present embodiment, a mixed solution of hydrochloric acid, cupric chloride, ferric chloride, and sodium chloride can be used as the leachate used in the pretreatment step.
酸性水溶液の具体例としては、塩化物イオンを１００〜２００ｇ／Ｌ、銅イオンを１〜３０ｇ／Ｌ、鉄イオンを１〜１０ｇ／Ｌ含有する５０〜１００℃の酸性水溶液を用いることが好ましい。すなわち、浸出工程における浸出液として酸性塩化浴を使用することで、モリブデン精鉱中の銅だけでなくカルシウムをもより効率的に除去できる。その結果、モリブデン精鉱の酸化焙焼処理で得られる酸化レニウムの揮発率を向上させることができる。更に銅イオンを浸出液中に存在させておくことで、銅の浸出反応も促進される。 As a specific example of the acidic aqueous solution, it is preferable to use an acidic aqueous solution at 50 to 100 ° C. containing 100 to 200 g / L of chloride ions, 1 to 30 g / L of copper ions, and 1 to 10 g / L of iron ions. That is, by using an acidic chloride bath as a leaching solution in the leaching step, not only copper in molybdenum concentrate but also calcium can be more efficiently removed. As a result, the volatilization rate of rhenium oxide obtained by oxidation roasting treatment of molybdenum concentrate can be improved. Furthermore, the copper leaching reaction is also promoted by allowing copper ions to be present in the leaching solution.
浸出液中の塩化物イオンの濃度は、銅及びカルシウムの溶解反応を高い効率で実現する観点から、１００ｇ／Ｌ以上であることが好ましく、１２０ｇ／Ｌ以上であることがより好ましく、１４０ｇ／Ｌ以上であることが更により好ましい。しかしながら、経済性を考慮すると、過度に高濃度にする必要はなく、浸出液中の塩化物イオンの濃度は一般には２００ｇ／Ｌ以下であり、好ましくは１８０ｇ／Ｌ以下である。 The concentration of chloride ions in the leachate is preferably 100 g / L or more, more preferably 120 g / L or more, and 140 g / L or more from the viewpoint of realizing a copper and calcium dissolution reaction with high efficiency. Even more preferably. However, in consideration of economy, it is not necessary to make the concentration too high, and the concentration of chloride ions in the leachate is generally 200 g / L or less, preferably 180 g / L or less.
浸出液中の銅イオンの濃度は、銅及びカルシウム浸出反応の促進の観点から、１ｇ／Ｌ以上であることが好ましく、５ｇ／Ｌ以上であることがより好ましい。しかしながら、経済性を考慮すると、過度に高濃度にする必要はなく、浸出液中の銅イオンの濃度は一般には３０ｇ／Ｌ以下であり、好ましくは２０ｇ／Ｌ以下である。 The concentration of copper ions in the leaching solution is preferably 1 g / L or more, and more preferably 5 g / L or more, from the viewpoint of promoting the copper and calcium leaching reaction. However, in consideration of economy, it is not necessary to make the concentration too high, and the concentration of copper ions in the leachate is generally 30 g / L or less, preferably 20 g / L or less.
鉄イオンは銅及びカルシウム浸出の促進に好適な成分であり、銅の溶解反応を高い効率で実現する観点から、１ｇ／Ｌ以上であることが好ましいが、２０ｇ／Ｌを超えるとＭｏの浸出率が顕著に増加して逸損することから、浸出液中の鉄イオン濃度は２０ｇ／Ｌ以下とするべきであり、１０ｇ／Ｌ以下であることが好ましく、８ｇ／Ｌ以下であることがより好ましく、６ｇ／Ｌ以下とすることがさらに好ましい。 Iron ion is a suitable component for promoting copper and calcium leaching, and is preferably 1 g / L or more from the viewpoint of realizing a copper dissolution reaction with high efficiency, but if it exceeds 20 g / L, the leaching rate of Mo Iron concentration in the leachate should be 20 g / L or less, preferably 10 g / L or less, more preferably 8 g / L or less, and 6 g / L or less is more preferable.
Ｍｏの浸出を防止するという観点からは、浸出液中の銅イオンと鉄イオンの合計が２５ｇ／Ｌ以下であることが好ましく、２０ｇ／Ｌ以下であることがより好ましい。 From the viewpoint of preventing leaching of Mo, the total of copper ions and iron ions in the leaching solution is preferably 25 g / L or less, and more preferably 20 g / L or less.
なお、上記の塩化物イオン、銅イオン及び鉄イオンの濃度は、酸性水溶液をモリブデン精鉱に接触させる前の浸出液中の濃度を指す。 In addition, the density | concentration of said chloride ion, copper ion, and iron ion points out the density | concentration in the leaching solution before making acidic aqueous solution contact molybdenum concentrate.
浸出液とモリブデン精鉱の接触方法としては特に制限はなく、噴霧や浸漬などの方法があるが、反応効率の観点から、浸出液中にモリブデン精鉱を浸漬し、撹拌する方法が好ましい。 The contact method between the leachate and the molybdenum concentrate is not particularly limited, and there are methods such as spraying and dipping. From the viewpoint of reaction efficiency, a method of dipping and stirring the molybdenum concentrate in the leachate is preferable.
上記浸出処理においては、酸素含有気体を浸出液に供給することにより銅及びカルシウムの浸出速度を高めることが可能となる。酸素含有気体の流量を増大させることで、銅及びカルシウムの浸出速度が増大する傾向にある。これにより、モリブデンが浸出するよりも先に銅及びカルシウムの浸出が進行するため、モリブデンの逸損を抑えることが可能となる。酸素含有気体としては、特に制限はないが、例えば空気、酸素、酸素と不活性ガス（窒素や希ガスなど）の混合ガスが挙げられる。経済性の観点からは空気が好ましい。 In the leaching process, it is possible to increase the leaching rate of copper and calcium by supplying an oxygen-containing gas to the leaching solution. By increasing the flow rate of the oxygen-containing gas, the leaching rate of copper and calcium tends to increase. Thereby, since the leaching of copper and calcium proceeds before leaching of molybdenum, it is possible to suppress the loss of molybdenum. Although there is no restriction | limiting in particular as oxygen-containing gas, For example, the mixed gas of air, oxygen, oxygen, and inert gas (nitrogen, a noble gas, etc.) is mentioned. Air is preferable from the viewpoint of economy.
酸素含有気体は上述した効果を有効に発揮させるという観点から前記浸出液１Ｌ当たり０．０２ｓｌｐｍ以上の流量で供給することが好ましく、０．０４ｓｌｐｍ以上の流量で供給することがより好ましく、０．０８ｓｌｐｍ以上の流量で供給することが更により好ましい。ただし、過剰に供給した場合は、気泡中への液の蒸発で奪われる蒸発熱を補償するために電力などのエネルギーを多く消費し、また、精鉱粒子が表面に塗された気泡の層（フロス）が大量に発生して反応槽からあふれるため、前記浸出液１Ｌ当たり０．５ｓｌｐｍ以下の流量で供給することが好ましく、前記浸出液１Ｌ当たり０．２５ｓｌｐｍ以下の流量で供給することがより好ましく、前記浸出液１Ｌ当たり０．１５ｓｌｐｍ以下の流量で供給することが更により好ましい。 The oxygen-containing gas is preferably supplied at a flow rate of 0.02 slpm or more, more preferably 0.04 slpm or more, and 0.08 slpm or more from the viewpoint of effectively exhibiting the above-described effects. It is even more preferable to supply at a flow rate of However, if it is supplied excessively, it consumes a lot of energy such as electric power to compensate for the heat of evaporation taken away by evaporation of the liquid into the bubbles, and a layer of bubbles with concentrate particles coated on the surface ( Is generated at a flow rate of 0.5 slpm or less per liter of the leachate, more preferably at a flow rate of 0.25 slpm or less per liter of the leachate, It is even more preferable to supply at a flow rate of 0.15 slpm or less per liter of leachate.
或いは、浸出工程は、浸出終点における浸出液の酸化還元電位（ｖｓ Ａｇ／ＡｇＣｌ）が４００〜４８０ｍＶの範囲に入るようにして行うこともまた、Ｍｏの溶出を抑えながらＣｕ及びＣａをより効率的に浸出させる観点で有効である。酸化還元電位が４８０ｍＶを超えるとＭｏの溶出が無視できなくなる一方で、酸化還元電位が４００ｍＶ未満だと銅及びカルシウムの浸出速度が極端に遅くなってしまい工業的ではない。浸出終点における浸出液の酸化還元電位は好ましくは４６０ｍＶ以下であり、より好ましくは４５０ｍＶ以下であり、更により好ましくは４４０ｍＶ以下であり、更に好ましくは４３０ｍＶであり、最も好ましくは４２０ｍＶ以下である。 Alternatively, the leaching step can be performed such that the redox potential (vs Ag / AgCl) of the leachate at the leaching end point is in the range of 400 to 480 mV. Effective from the viewpoint of leaching. When the redox potential exceeds 480 mV, the elution of Mo cannot be ignored. On the other hand, when the redox potential is less than 400 mV, the leaching rate of copper and calcium becomes extremely slow, which is not industrial. The redox potential of the leachate at the end of leaching is preferably 460 mV or less, more preferably 450 mV or less, even more preferably 440 mV or less, still more preferably 430 mV, and most preferably 420 mV or less.
浸出終点における浸出液の酸化還元電位は好ましくは４１０ｍＶ以上である。浸出液の酸化還元電位は、酸素含有ガスを吹き込む等の特別な操作を行わない限り浸出時間の経過に伴って徐々に低下することから、特別な操作を行わない場合に浸出終点における酸化還元電位をこのような範囲に制御するためには、浸出初期段階で酸化還元電位が十分に高くなるように液組成及びパルプ濃度を設定することが重要となる。 The redox potential of the leaching solution at the leaching end point is preferably 410 mV or more. The oxidation-reduction potential of the leachate gradually decreases with the lapse of the leaching time unless a special operation such as blowing in oxygen-containing gas is performed. In order to control within such a range, it is important to set the liquid composition and the pulp concentration so that the oxidation-reduction potential is sufficiently high at the initial stage of leaching.
酸化還元電位は浸出終点のみならず、浸出最中においても上記の範囲に維持されることが望ましい。安定した浸出効果を得るためである。具体的には、浸出開始時点から２時間経過後から浸出終点まで上述した範囲の酸化還元電位に維持することが好ましく、浸出開始時点から１時間経過後から浸出終点まで上述した範囲の酸化還元電位に維持することがより好ましい。 It is desirable that the oxidation-reduction potential be maintained in the above range not only during the leaching end point but also during leaching. This is to obtain a stable leaching effect. Specifically, it is preferable to maintain the redox potential in the above-described range from 2 hours after the start of leaching until the end of leaching, and the redox potential in the above range from 1 hour after the start of leaching to the end of leaching. It is more preferable to maintain it.
浸出開始時点というのは浸出液とモリブデン精鉱の接触が開始された時点であり、浸出終点というのは浸出液をモリブデン精鉱から固液分離した時点をいう。 The leaching start time is the time when contact between the leachate and the molybdenum concentrate starts, and the leaching end point means the time when the leachate is solid-liquid separated from the molybdenum concentrate.
前処理工程に使用する浸出液の温度は浸出効率や装置の材質の観点から、５０℃以上とするのが好ましく、６０℃以上とするのがより好ましく、７０℃以上とするのが更により好ましいが、高すぎると浸出液の蒸発や加熱コストの上昇あるので、１００℃以下とするのが好ましく、９０℃以下とするのがより好ましく、８５℃以下とするのがより好ましい。浸出効率を高めることを目的として前処理工程を加圧下で実施することも可能であるが、大気圧下で十分である。すなわち高圧での浸出工程を行うための耐圧容器を必要とせず、より簡易な装置を用いることができる。銅及びカルシウム浸出を促進するため、処理対象となるモリブデン精鉱を予め粉砕・摩鉱しておくことが好ましい。 The temperature of the leachate used in the pretreatment step is preferably 50 ° C. or higher, more preferably 60 ° C. or higher, and even more preferably 70 ° C. or higher, from the viewpoint of leaching efficiency and material of the apparatus. If it is too high, the leachate will evaporate and the heating cost will increase. Therefore, it is preferably 100 ° C. or lower, more preferably 90 ° C. or lower, and more preferably 85 ° C. or lower. It is possible to carry out the pretreatment step under pressure for the purpose of increasing the leaching efficiency, but it is sufficient under atmospheric pressure. That is, a pressure vessel for performing a high pressure leaching step is not required, and a simpler apparatus can be used. In order to promote leaching of copper and calcium, it is preferable to pulverize and grind the molybdenum concentrate to be treated in advance.
前処理工程では、使用する浸出液に対するモリブデン精鉱の量を大きくして実施する方が浸出コストの低減の観点から好ましい。そのため、本発明の一実施形態においては、５０ｇ／Ｌ以上のパルプ濃度で浸出工程を行うことができ、本発明の別の一実施形態においては、１５０ｇ／Ｌ以上のパルプ濃度で浸出工程を行うことができ、本発明の更に別の一実施形態においては、３００ｇ／Ｌ以上のパルプ濃度で浸出工程を行うことができる。一方で、浸出速度を高めるという観点からは使用する浸出液に対するモリブデン精鉱の量は小さい方が好ましいことから、本発明の一実施形態においては、８００ｇ／Ｌ以下のパルプ濃度で浸出工程を行うことができ、本発明の別の一実施形態においては、６００ｇ／Ｌ以下のパルプ濃度で浸出工程を行うことができ、本発明の更に別の一実施形態においては、５００ｇ／Ｌ以下のパルプ濃度で浸出工程を行うことができる。ここで、パルプ濃度とは使用する浸出液の体積（Ｌ）に対するモリブデン精鉱（乾燥重量（ｇ））の比である。 In the pretreatment step, it is preferable to increase the amount of molybdenum concentrate with respect to the leaching solution used from the viewpoint of reducing the leaching cost. Therefore, in one embodiment of the present invention, the leaching step can be performed at a pulp concentration of 50 g / L or more, and in another embodiment of the present invention, the leaching step is performed at a pulp concentration of 150 g / L or more. In yet another embodiment of the present invention, the leaching step can be performed at a pulp concentration of 300 g / L or more. On the other hand, from the viewpoint of increasing the leaching rate, it is preferable that the amount of molybdenum concentrate with respect to the leaching solution to be used is smaller. In another embodiment of the present invention, the leaching step can be performed at a pulp concentration of 600 g / L or less, and in yet another embodiment of the present invention, a pulp concentration of 500 g / L or less. A leaching process can be performed. Here, the pulp concentration is the ratio of molybdenum concentrate (dry weight (g)) to the volume (L) of the leachate used.
銅及びカルシウムの浸出工程は一段階で実施することもできるが、モリブデン精鉱中の銅の浸出を十分に行うために浸出工程を複数段階で実施することも可能である。複数段階を利用した浸出工程は、具体的には、一段目における浸出操作を終了後に、フィルタープレスやシックナーなどによって固液分離し、浸出残渣に対して次段の銅浸出操作を行うことにより実施することができる。 The leaching process of copper and calcium can be carried out in one stage, but the leaching process can also be carried out in a plurality of stages in order to sufficiently brew copper in the molybdenum concentrate. Specifically, the leaching process using multiple stages is carried out by performing solid-liquid separation with a filter press or thickener after completing the leaching operation in the first stage and performing the next stage copper leaching operation on the leaching residue. can do.
本発明によれば、モリブデンの浸出を抑制しながらも、銅及びカルシウムを浸出することが可能である。例えば、本発明の一実施形態においては、浸出後液中のモリブデン濃度を０．００１ｇ／Ｌ以下に抑制しながら、銅の浸出率７０％以上を達成することができる。本発明の別の一実施形態においては、浸出後液中のモリブデン濃度を０．００５ｇ／Ｌ以下に抑制しながら、銅の浸出率９０％以上を達成することができ、本発明の更に別の一実施形態においては、浸出後液中のモリブデン濃度を０．００５ｇ／Ｌ以下に抑制しながら、銅の浸出率９５％以上を達成することができる。 According to the present invention, copper and calcium can be leached while suppressing leaching of molybdenum. For example, in one embodiment of the present invention, a copper leaching rate of 70% or more can be achieved while suppressing the molybdenum concentration in the liquid after leaching to 0.001 g / L or less. In another embodiment of the present invention, it is possible to achieve a copper leaching rate of 90% or more while suppressing the molybdenum concentration in the liquid after leaching to 0.005 g / L or less. In one embodiment, a copper leaching rate of 95% or more can be achieved while suppressing the molybdenum concentration in the liquid after leaching to 0.005 g / L or less.
また、浸出工程は、浸出残渣中の銅品位が２．０質量％以下、好ましくは１．５質量％以下とし、典型的には０．３〜１．２質量％程度になるまで実施することが好ましい。これにより、後述する酸化焙焼工程におけるＲｅの揮発率を８５％以上に高めることができる。 In addition, the leaching step is carried out until the copper grade in the leaching residue is 2.0 mass% or less, preferably 1.5 mass% or less, and typically about 0.3 to 1.2 mass%. Is preferred. Thereby, the volatilization rate of Re in the oxidation roasting step described later can be increased to 85% or more.
或いは、浸出工程は、浸出残渣中のカルシウム品位が１５００ｐｐｍ以下、好ましくは５００ｐｐｍ以下になるまで実施することが好ましい。なお、カルシウム品位の下限は特に制限されないが、浸出処理時間と生産性との観点から２００ｐｐｍ以上とすることが好ましい。これにより、後述する酸化焙焼工程におけるＲｅの揮発率を８０％以上、より典型的には８５％以上に高めることができる。 Alternatively, the leaching step is preferably carried out until the calcium quality in the leaching residue is 1500 ppm or less, preferably 500 ppm or less. The lower limit of calcium quality is not particularly limited, but is preferably 200 ppm or more from the viewpoint of leaching treatment time and productivity. Thereby, the volatility of Re in the oxidation roasting step described later can be increased to 80% or more, more typically 85% or more.
或いは、浸出工程は、浸出残渣中の鉄品位が１０質量％以下、好ましくは５質量％以下、より好ましくは３．６質量％以下、典型的には２．０〜３．８質量％、より好ましくは２．５〜３．５質量％になるまで実施することが好ましい。 Alternatively, in the leaching step, the iron grade in the leaching residue is 10% by mass or less, preferably 5% by mass or less, more preferably 3.6% by mass or less, typically 2.0 to 3.8% by mass, or more. It is preferable to carry out until 2.5 to 3.5% by mass.
浸出に要する時間は、原料であるモリブデン精鉱中の銅品位にもよるが、浸出残渣中の銅品位が０．５質量％となるまでに要する時間は、例えば４〜１０時間くらいであり、典型的には５〜６時間くらいである。 The time required for the leaching depends on the copper grade in the molybdenum concentrate as a raw material, but the time required for the copper grade in the leaching residue to be 0.5 mass% is, for example, about 4 to 10 hours, Typically about 5-6 hours.
（２）酸化焙焼工程
次に、前処理工程で固液分離により得られた浸出残渣（モリブデン精鉱処理物）を酸化焙焼する。酸化焙焼処理では、処理物の処理温度を５００〜６００℃、より典型的には５５０℃前後とすることが好ましい。酸化焙焼処理には、空気などの酸素含有気体を焙焼炉に供給する。典型的には２．０〜８．０（Ｌ／ｇ−ＭｏＳ2）、より典型的には、２．５〜６．０（Ｌ／ｇ−ＭｏＳ2）となるように酸素含有気体を供給することが好ましい。焙焼時間は特に制限されないが、例えば１．５〜６時間程度とすることができる。
(2) Oxidation roasting step Next, the leaching residue (molybdenum concentrate processed product) obtained by solid-liquid separation in the pretreatment step is oxidized and roasted. In the oxidative roasting treatment, the treatment temperature of the treated product is preferably 500 to 600 ° C, more typically around 550 ° C. For the oxidation roasting treatment, an oxygen-containing gas such as air is supplied to the roasting furnace. The oxygen-containing gas is supplied so as to be typically 2.0 to 8.0 (L / g-MoS 2 ), more typically 2.5 to 6.0 (L / g-MoS 2 ). It is preferable to do. The roasting time is not particularly limited, but can be, for example, about 1.5 to 6 hours.
（３）酸化レニウム回収工程
焙焼工程で揮発した酸化レニウムは、ダストとして、もしくは排ガス洗浄における洗浄水に溶解し、溶解液（以下溶解液を「廃酸」とも表記する）として回収される。
(3) Rhenium oxide recovery step The rhenium oxide volatilized in the roasting step is dissolved as dust or in washing water in exhaust gas cleaning, and recovered as a solution (hereinafter, the solution is also referred to as “waste acid”).
（レニウム回収）
ダストもしくは廃酸からのレニウムの回収は周知の技術にて回収できる。例えば、ダストを水や酸により溶解し得た溶解液、もしくは廃酸を強塩基性樹脂によりＲｅを分離し、その後アンモニア性溶液にて溶離し過レニウム酸アンモニウム（ＡＰＲ）として回収できる。
(Rhenium recovery)
Recovery of rhenium from dust or spent acid can be recovered by well-known techniques. For example, it is possible to recover Re as ammonium perrhenate (APR) by separating Re from a solution obtained by dissolving dust with water or acid, or by separating waste acid from a strongly basic resin and then eluting with an ammoniacal solution.
＜その他＞
（鉄酸化）
浸出工程で得られた浸出液には、浸出液に元々含まれていた鉄の他、モリブデン精鉱中の鉄の一部が溶解した鉄が含まれている。これらの鉄イオンの多くはＦｅ(II)と考えられる。これを酸化することでＦｅ(III)とし、再度浸出に使用することができる。またｐＨを調整することで、Ｆｅ（III）の一部を沈殿し、浸出液中の鉄濃度をコントロールすることができる。酸化の条件としては２０〜７０℃、ｐＨ１．５〜３．０でエアレーションすることがコストと反応速度の面で最も好ましい。温度とｐＨは高ければ高いほど、反応速度が速い。
<Others>
(Iron oxidation)
The leachate obtained in the leaching step contains iron in which a part of iron in the molybdenum concentrate is dissolved in addition to iron originally contained in the leachate. Many of these iron ions are thought to be Fe (II). This can be oxidized to Fe (III) and used again for leaching. Further, by adjusting the pH, a part of Fe (III) is precipitated, and the iron concentration in the leachate can be controlled. As oxidation conditions, aeration at 20 to 70 ° C. and pH 1.5 to 3.0 is most preferable in terms of cost and reaction rate. The higher the temperature and pH, the faster the reaction rate.
（銅回収）
浸出工程で得られた浸出後液から銅を回収することができる。銅の回収方法としては特に制限はないが、例えば溶媒抽出、イオン交換、卑な金属との置換析出及び電解採取などを利用することができる。浸出後液中の銅は１価及び２価の状態が混在しているが、溶媒抽出やイオン交換を円滑に行うために、全部が２価の銅イオンとなるように予め酸化しておくことが好ましい。酸化の方法は特に制限はないが空気や酸素を浸出後液中に吹き込む方法が簡便である。
(Copper recovery)
Copper can be recovered from the post-leaching solution obtained in the leaching step. Although there is no restriction | limiting in particular as a copper collection | recovery method, For example, solvent extraction, ion exchange, substitution precipitation with a base metal, electrowinning, etc. can be utilized. The copper in the solution after leaching contains both monovalent and divalent states, but in order to perform solvent extraction and ion exchange smoothly, all of them should be oxidized beforehand to be divalent copper ions. Is preferred. The method of oxidation is not particularly limited, but a method of leaching air or oxygen into the liquid after leaching is simple.
本発明に係るモリブデン精鉱の処理方法の一実施形態においては、浸出後液中の銅を、溶媒抽出及び逆抽出を経て、電解採取によりカソード上に電気銅として回収する工程を更に含む。当該工程自体は一般にＳＸ−ＥＷ（Solvent Extraction and Electro-Winning）法と呼ばれている方法であり、当業者にとって周知である。 In one embodiment of the method for treating molybdenum concentrate according to the present invention, the method further includes the step of recovering copper in the solution after leaching as electrolytic copper on the cathode by electrolytic extraction through solvent extraction and back extraction. The process itself is generally called a SX-EW (Solvent Extraction and Electro-Winning) method and is well known to those skilled in the art.
また、溶媒抽出前に、浸出後液に空気などの酸素含有気体を吹き込んで液中の銅を酸化する工程を経ることもできる。これにより、銅を溶媒抽出後に逆抽出（ストリップ）して直接電解採取することを可能にするという利点が得られる。酸化工程を経ない場合、強塩化物浴では一価の銅が高濃度で存在するため電解採取の際にデンドライト銅として析出する。デンドライト銅は金属粉末として電解槽に沈殿する。カソードに板状銅として回収する方が圧倒的に運搬等の操作性の面で長所が多い。更に、酸化工程後は固液分離することもできる。固液分離は浸出液中に鉄が含まれている場合に、酸化残渣に移行するため、浸出
後液中の銅純度を高める上で有利である。
In addition, before the solvent extraction, a step of oxidizing copper in the liquid by blowing an oxygen-containing gas such as air into the liquid after leaching can also be performed. This provides the advantage that copper can be back extracted (striped) after solvent extraction and directly electrowinned. In the case of not passing through the oxidation step, monovalent copper is present in a high concentration in the strong chloride bath, so that it is deposited as dendrite copper during electrowinning. Dendritic copper precipitates in the electrolytic cell as a metal powder. It is overwhelmingly more advantageous in terms of operability such as transportation when recovered as plate-like copper on the cathode. Furthermore, solid-liquid separation can also be performed after the oxidation step. Solid-liquid separation is advantageous in increasing the copper purity in the liquid after leaching because it moves to an oxidation residue when iron is contained in the leaching liquid.
以下に本発明の実施例を比較例と共に示すが、これらの実施例は本発明及びその利点をよりよく理解するために提供するものであり、発明が限定されることを意図するものではない。 Examples of the present invention will be described below together with comparative examples, but these examples are provided for better understanding of the present invention and its advantages, and are not intended to limit the invention.
（実施例１：前処理の有無による揮発率向上の影響）
輝水鉛鉱を含む鉱石から浮遊選鉱により選別されたモリブデン精鉱を粉砕したものを用意した。モリブデン精鉱は、酸溶解後にＩＣＰ発光分光分析法（ＩＣＰ−ＯＥＳ）で分析したところ、Ｍｏ：４５質量％、Ｃｕ：３．６質量％、Ｆｅ：４．５質量％、Ｓ：３７質量％、Ｒｅ：３５０ｐｐｍ、Ｃａ：３７００ｐｐｍの組成を有していた。全塩化物イオン１８０ｇ／Ｌ、Ｃｕイオン１８ｇ／Ｌ、Ｆｅイオン２ｇ／Ｌ、塩酸７ｇ／Ｌのイオン組成を有し、ｐＨが０．７５となるように、塩酸、塩化第二鉄、塩化第二銅、塩化ナトリウムを混合した浸出液（酸性水溶液）を調整し、この浸出液４Ｌをホットスターラーで７５℃に加熱後、当該モリブデン精鉱１５２０ｇを投入し、浸出液への空気吹き込み（０．３７ｓｌｐｍ）と撹拌を継続しながら浸出試験を実施し、固液分離を行って浸出後液及び浸出残渣を得た。なお、浸出液及び残渣中の金属の分析は、ＩＣＰ発光分光分析法で行った。イオン濃度は、浸出液の成分が完全に電離していると仮定して算出した。
(Example 1: Influence of volatilization rate improvement by presence or absence of pre-processing)
We prepared a pulverized molybdenum concentrate selected by floatation from ores containing molybdenite. Molybdenum concentrate was analyzed by ICP emission spectroscopy (ICP-OES) after acid dissolution, and was found to have Mo: 45 mass%, Cu: 3.6 mass%, Fe: 4.5 mass%, S: 37 mass%. , Re: 350 ppm, Ca: 3700 ppm. It has an ionic composition of total chloride ions of 180 g / L, Cu ions of 18 g / L, Fe ions of 2 g / L, and hydrochloric acid of 7 g / L, so that the pH is 0.75. After preparing a leachate (acidic aqueous solution) in which dicopper and sodium chloride were mixed, 4L of this leachate was heated to 75 ° C. with a hot stirrer, and then 1520 g of the molybdenum concentrate was added and air blowing into the leachate (0.37 slpm) A leaching test was carried out while continuing stirring, and solid-liquid separation was performed to obtain a liquid after leaching and a leaching residue. The metal in the leachate and the residue was analyzed by ICP emission spectroscopic analysis. The ion concentration was calculated on the assumption that the components of the leachate were completely ionized.
固液分離によって得られた浸出残渣をＩＣＰ発光分光分析法で分析したところ、Ｍｏ：５１質量％、Ｃｕ：０．４５〜０．４７質量％、Ｆｅ：２．９〜３．０質量％、Ｓ：４０質量％、Ｒｅ：３８０ｐｐｍ、Ｓｉ：１５０００ｐｐｍ、Ｃａ：１００ｐｐｍ未満〜１０００ｐｐｍであった。表１に浸出残渣の組成を示す（表１中、従来例１〜３は浸出処理を行わないモリブデン精鉱（生鉱）の組成を示す）。 When the leaching residue obtained by solid-liquid separation was analyzed by ICP emission spectrometry, Mo: 51% by mass, Cu: 0.45-0.47% by mass, Fe: 2.9-3.0% by mass, S: 40% by mass, Re: 380 ppm, Si: 15000 ppm, Ca: less than 100 ppm to 1000 ppm. The composition of the leaching residue is shown in Table 1 (in Table 1, Conventional Examples 1 to 3 show the composition of molybdenum concentrate (raw ore) not subjected to leaching treatment).
表１に示す浸出残渣に対して、表２に示す条件で焙焼時間とガス速度を変化させながら酸化焙焼処理を行った。各焙焼時間に対する酸化焙焼処理で得られた酸化レニウムの揮発率（Ｒｅ揮発率［％］）の関係を図１に示す。なお、比較のため、上記の前処理（浸出・固液分離）を行わないモリブデン精鉱も同様に、焙焼時間とガス速度を変化させながら酸化焙焼処理を行ってＲｅ揮発率を比較した。Ｒｅ揮発率と酸化焙焼処理で得られた焙焼残渣の残渣率の評価結果を表２に示す。 The leaching residue shown in Table 1 was subjected to oxidation roasting treatment while changing the roasting time and gas velocity under the conditions shown in Table 2. The relationship of the volatilization rate (Re volatilization rate [%]) of rhenium oxide obtained by the oxidation roasting process with respect to each roasting time is shown in FIG. For comparison, molybdenum concentrates not subjected to the above pretreatment (leaching / solid-liquid separation) were similarly subjected to oxidation roasting treatment while changing the roasting time and gas velocity, and the Re volatility was compared. . Table 2 shows the evaluation results of the Re volatilization rate and the residue rate of the roasting residue obtained by the oxidation roasting treatment.
Ｒｅ揮発率の評価は、焙焼時にＳｉが揮発しないものとし、以下の式（１）、（２）に示す計算式にて算出した。
Ｒｅ揮発率（％）＝１００×（１−（残渣Ｒｅ品位（ｐｐｍ）×残渣率（％）／原料Ｒｅ品位（ｐｐｍ）×１００））・・・（１）
残渣率（％）＝１００×（原料Ｓｉ品位（ｐｐｍ）／残渣Ｓｉ品位（ｐｐｍ））・・・（２）
The evaluation of the Re volatility was calculated according to the following formulas (1) and (2) assuming that Si does not volatilize during roasting.
Re volatilization rate (%) = 100 × (1- (residue Re quality (ppm) × residue rate (%) / raw material Re quality (ppm) × 100)) (1)
Residual rate (%) = 100 × (raw material Si quality (ppm) / residual Si quality (ppm)) (2)
図１に示すように、焙焼処理前に、本発明に係るモリブデン精鉱から銅及びカルシウムを除去するための前処理を行うことによって（本発明例１〜３）、前処理を行わない従来例１〜３に比べていずれも１０％程度、レニウム揮発率が向上していることが分かる。 As shown in FIG. 1, prior to the roasting treatment, by performing a pretreatment for removing copper and calcium from the molybdenum concentrate according to the present invention (Examples 1 to 3 of the present invention), no conventional pretreatment is performed. It can be seen that the rhenium volatility is improved by about 10% compared to Examples 1 to 3.
（実施例２：前処理で低減されるＣｕ、Ｃａ、Ｆｅ品位に対するＲｅ揮発率の影響）
実施例１と同様のモリブデン精鉱に対して、酸性浸出液１として実施例１と同様の浸出液を用いて、酸性浸出液２としてモリブデン精鉱中のＣａを除去するために塩酸を浸出液として用いて浸出処理を行い、固液分離後、表３に示す組成の浸出残渣を得た。浸出残渣に対し、表４に示す条件で酸化焙焼処理を行って酸化レニウムを回収し、Ｒｅ揮発率を算出した。比較のため、上記の前処理（浸出・固液分離）を行わないモリブデン精鉱も同様に酸化焙焼処理を行って酸化レニウムを回収し、Ｒｅ揮発率と焙焼残渣の残渣率を算出した。
(Example 2: Effect of Re volatilization rate on Cu, Ca, Fe quality reduced by pretreatment)
Leaching the same molybdenum concentrate as in Example 1 using the same leachate as in Example 1 as acidic leachate 1 and using hydrochloric acid as the leachate to remove Ca in the molybdenum concentrate as acidic leachate 2 After the treatment and solid-liquid separation, a leaching residue having the composition shown in Table 3 was obtained. The leaching residue was subjected to oxidation roasting treatment under the conditions shown in Table 4 to recover rhenium oxide, and the Re volatilization rate was calculated. For comparison, molybdenum concentrates not subjected to the above pretreatment (leaching / solid-liquid separation) were also subjected to oxidation roasting treatment to recover rhenium oxide, and the Re volatilization rate and the residue rate of the roasting residue were calculated. .
図２に示すように、前処理を行わないモリブデン精鉱を酸化焙焼処理した場合にはＲｅ揮発率が８０％を下回っていたのに対し、酸性浸出液２によって脱Ｃａ処理した場合はＲｅ揮発率が８０％を超えた。更に、酸性浸出液１によって脱銅・脱Ｃａ処理した場合には、Ｒｅ揮発率を最大で８９．６％まで向上させることができた。また、図２〜図４より、酸化焙焼対象処理物のＣｕ、Ｆｅ、Ｃａ品位を調整することでＲｅ揮発率が向上できることがわかる。これは、前処理後のモリブデン精鉱処理物が、前処理前のモリブデン精鉱に比べてＣａ、Ｃｕ品位が低減されているため、酸化焙焼処理において、処理物中のＣａ、Ｃｕと反応してＣａ（ＲｅＯ4）2、Ｃｕ（ＲｅＯ4）2等の形態で揮発しなかったためと考えられる。 As shown in FIG. 2, the Re volatilization rate was lower than 80% when the molybdenum concentrate without pretreatment was oxidized and roasted, whereas the Re volatilization was performed when the Ca was removed by the acidic leachate 2. The rate exceeded 80%. Furthermore, when the copper removal and the Ca removal treatment were performed with the acidic leachate 1, the Re volatilization rate could be improved up to 89.6%. Moreover, it turns out that Re volatilization rate can be improved from FIGS. 2-4 by adjusting Cu, Fe, and Ca quality of the oxidation roasting object processed material. This is because the treated product of molybdenum concentrate after pre-treatment has reduced Ca and Cu quality compared to the molybdenum concentrate before pre-treatment, so that it reacts with Ca and Cu in the treated product in the oxidation roasting treatment. This is considered to be because it was not volatilized in the form of Ca (ReO 4 ) 2 , Cu (ReO 4 ) 2 or the like.
また、本実施形態に係る前処理を行ったモリブデン精鉱を酸化焙焼することによって、焙焼後に得られる残渣中の残渣Ｓ品位をも低減することができる。これにより、焙焼処理残渣中の硫黄濃度を低減することができるため、回収される三酸化モリブデンの純度をより向上可能であることも分かる。 Moreover, the residue S quality in the residue obtained after roasting can also be reduced by carrying out oxidation roasting of the molybdenum concentrate which performed the pre-processing which concerns on this embodiment. Thereby, since the sulfur density | concentration in a baking process residue can be reduced, it turns out that the purity of the collect | recovered molybdenum trioxide can be improved more.
Claims (3)
前記前処理工程で得られたモリブデン精鉱処理物を酸化焙焼する酸化焙焼工程と、
前記酸化焙焼工程において揮発した酸化レニウムを回収する工程と
を含むモリブデン精鉱からのレニウムの回収方法。 Contact the molybdenum concentrate while supplying oxygen-containing gas to an acidic aqueous solution at 50 to 100 ° C. containing 100 to 200 g / L of chloride ions, 1 to 30 g / L of copper ions, and 1 to 10 g / L of iron ions. A pretreatment step including a step of leaching copper and calcium contained in molybdenum concentrate and a step of obtaining a liquid and a leaching residue after leaching by solid-liquid separation ;
An oxidation roasting step of oxidizing and roasting the molybdenum concentrate processed product obtained in the pretreatment step;
A method for recovering rhenium from molybdenum concentrate, the method including recovering volatilized rhenium oxide in the oxidation roasting step.
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