CN115323194B - Process method for treating molybdenite by full wet method under normal pressure condition - Google Patents
Process method for treating molybdenite by full wet method under normal pressure condition Download PDFInfo
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- 238000000034 method Methods 0.000 title claims abstract description 82
- 230000008569 process Effects 0.000 title claims abstract description 57
- CWQXQMHSOZUFJS-UHFFFAOYSA-N molybdenum disulfide Chemical compound S=[Mo]=S CWQXQMHSOZUFJS-UHFFFAOYSA-N 0.000 title claims abstract description 41
- 229910052961 molybdenite Inorganic materials 0.000 title claims abstract description 31
- QGZKDVFQNNGYKY-UHFFFAOYSA-N Ammonia Chemical compound N QGZKDVFQNNGYKY-UHFFFAOYSA-N 0.000 claims abstract description 98
- 238000002386 leaching Methods 0.000 claims abstract description 94
- 229910021529 ammonia Inorganic materials 0.000 claims abstract description 49
- 239000002253 acid Substances 0.000 claims abstract description 40
- 239000002893 slag Substances 0.000 claims abstract description 37
- 229910052740 iodine Inorganic materials 0.000 claims abstract description 36
- 239000011630 iodine Substances 0.000 claims abstract description 36
- JKQOBWVOAYFWKG-UHFFFAOYSA-N molybdenum trioxide Chemical compound O=[Mo](=O)=O JKQOBWVOAYFWKG-UHFFFAOYSA-N 0.000 claims abstract description 36
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 claims abstract description 35
- ZCYVEMRRCGMTRW-UHFFFAOYSA-N 7553-56-2 Chemical compound [I] ZCYVEMRRCGMTRW-UHFFFAOYSA-N 0.000 claims abstract description 29
- 239000012141 concentrate Substances 0.000 claims abstract description 27
- 229910052717 sulfur Inorganic materials 0.000 claims abstract description 25
- 238000005188 flotation Methods 0.000 claims abstract description 24
- 239000011593 sulfur Substances 0.000 claims abstract description 22
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims abstract description 20
- 229910052802 copper Inorganic materials 0.000 claims abstract description 20
- 239000010949 copper Substances 0.000 claims abstract description 20
- 230000003647 oxidation Effects 0.000 claims abstract description 20
- 238000007254 oxidation reaction Methods 0.000 claims abstract description 20
- 238000000926 separation method Methods 0.000 claims abstract description 18
- 229910052702 rhenium Inorganic materials 0.000 claims abstract description 15
- WUAPFZMCVAUBPE-UHFFFAOYSA-N rhenium atom Chemical compound [Re] WUAPFZMCVAUBPE-UHFFFAOYSA-N 0.000 claims abstract description 15
- UGFAIRIUMAVXCW-UHFFFAOYSA-N Carbon monoxide Chemical compound [O+]#[C-] UGFAIRIUMAVXCW-UHFFFAOYSA-N 0.000 claims abstract description 9
- 239000003546 flue gas Substances 0.000 claims abstract description 9
- 229910000510 noble metal Inorganic materials 0.000 claims abstract description 8
- 239000000126 substance Substances 0.000 claims abstract description 7
- 230000003197 catalytic effect Effects 0.000 claims abstract description 6
- 239000000243 solution Substances 0.000 claims description 49
- 239000007788 liquid Substances 0.000 claims description 35
- OKTJSMMVPCPJKN-UHFFFAOYSA-N Carbon Chemical compound [C] OKTJSMMVPCPJKN-UHFFFAOYSA-N 0.000 claims description 29
- 238000003756 stirring Methods 0.000 claims description 20
- 239000007787 solid Substances 0.000 claims description 19
- 229910000360 iron(III) sulfate Inorganic materials 0.000 claims description 18
- 239000011609 ammonium molybdate Substances 0.000 claims description 17
- 235000018660 ammonium molybdate Nutrition 0.000 claims description 17
- APUPEJJSWDHEBO-UHFFFAOYSA-P ammonium molybdate Chemical compound [NH4+].[NH4+].[O-][Mo]([O-])(=O)=O APUPEJJSWDHEBO-UHFFFAOYSA-P 0.000 claims description 17
- 229940010552 ammonium molybdate Drugs 0.000 claims description 17
- 238000000605 extraction Methods 0.000 claims description 17
- 238000005406 washing Methods 0.000 claims description 17
- RAHZWNYVWXNFOC-UHFFFAOYSA-N Sulphur dioxide Chemical compound O=S=O RAHZWNYVWXNFOC-UHFFFAOYSA-N 0.000 claims description 16
- 238000006243 chemical reaction Methods 0.000 claims description 16
- 235000021110 pickles Nutrition 0.000 claims description 16
- 239000000047 product Substances 0.000 claims description 16
- 230000002378 acidificating effect Effects 0.000 claims description 15
- -1 iodide ions Chemical class 0.000 claims description 14
- GEHJYWRUCIMESM-UHFFFAOYSA-L sodium sulfite Chemical compound [Na+].[Na+].[O-]S([O-])=O GEHJYWRUCIMESM-UHFFFAOYSA-L 0.000 claims description 13
- 239000011259 mixed solution Substances 0.000 claims description 11
- VHUUQVKOLVNVRT-UHFFFAOYSA-N Ammonium hydroxide Chemical compound [NH4+].[OH-] VHUUQVKOLVNVRT-UHFFFAOYSA-N 0.000 claims description 9
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 9
- 235000011114 ammonium hydroxide Nutrition 0.000 claims description 9
- 230000035484 reaction time Effects 0.000 claims description 9
- RUTXIHLAWFEWGM-UHFFFAOYSA-H iron(3+) sulfate Chemical compound [Fe+3].[Fe+3].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O RUTXIHLAWFEWGM-UHFFFAOYSA-H 0.000 claims description 8
- VTLYFUHAOXGGBS-UHFFFAOYSA-N Fe3+ Chemical compound [Fe+3] VTLYFUHAOXGGBS-UHFFFAOYSA-N 0.000 claims description 7
- UYJXRRSPUVSSMN-UHFFFAOYSA-P ammonium sulfide Chemical compound [NH4+].[NH4+].[S-2] UYJXRRSPUVSSMN-UHFFFAOYSA-P 0.000 claims description 7
- 238000007654 immersion Methods 0.000 claims description 7
- 238000002360 preparation method Methods 0.000 claims description 7
- 229910001447 ferric ion Inorganic materials 0.000 claims description 6
- 235000010265 sodium sulphite Nutrition 0.000 claims description 6
- 238000001704 evaporation Methods 0.000 claims description 5
- 241000605222 Acidithiobacillus ferrooxidans Species 0.000 claims description 4
- 230000009471 action Effects 0.000 claims description 4
- 238000007664 blowing Methods 0.000 claims description 4
- XMBWDFGMSWQBCA-UHFFFAOYSA-M iodide Chemical compound [I-] XMBWDFGMSWQBCA-UHFFFAOYSA-M 0.000 claims description 4
- 244000005700 microbiome Species 0.000 claims description 4
- 238000002156 mixing Methods 0.000 claims description 4
- 230000033116 oxidation-reduction process Effects 0.000 claims description 4
- QGZKDVFQNNGYKY-UHFFFAOYSA-O Ammonium Chemical compound [NH4+] QGZKDVFQNNGYKY-UHFFFAOYSA-O 0.000 claims description 3
- CWYNVVGOOAEACU-UHFFFAOYSA-N Fe2+ Chemical compound [Fe+2] CWYNVVGOOAEACU-UHFFFAOYSA-N 0.000 claims description 3
- 238000003795 desorption Methods 0.000 claims description 3
- 229910001448 ferrous ion Inorganic materials 0.000 claims description 3
- 239000007789 gas Substances 0.000 claims description 3
- 239000012535 impurity Substances 0.000 claims description 3
- 238000005554 pickling Methods 0.000 claims description 3
- HRLYFPKUYKFYJE-UHFFFAOYSA-N tetraoxorhenate(2-) Chemical compound [O-][Re]([O-])(=O)=O HRLYFPKUYKFYJE-UHFFFAOYSA-N 0.000 claims description 3
- 235000019738 Limestone Nutrition 0.000 claims description 2
- 238000002425 crystallisation Methods 0.000 claims description 2
- 230000008025 crystallization Effects 0.000 claims description 2
- 230000008020 evaporation Effects 0.000 claims description 2
- 239000006028 limestone Substances 0.000 claims description 2
- 230000001590 oxidative effect Effects 0.000 claims description 2
- 239000000843 powder Substances 0.000 claims description 2
- 238000001179 sorption measurement Methods 0.000 claims description 2
- 238000003860 storage Methods 0.000 claims description 2
- 239000006228 supernatant Substances 0.000 claims description 2
- 229910052750 molybdenum Inorganic materials 0.000 abstract description 39
- ZOKXTWBITQBERF-UHFFFAOYSA-N Molybdenum Chemical compound [Mo] ZOKXTWBITQBERF-UHFFFAOYSA-N 0.000 abstract description 31
- 239000011733 molybdenum Substances 0.000 abstract description 31
- 239000010931 gold Substances 0.000 abstract description 14
- PCHJSUWPFVWCPO-UHFFFAOYSA-N gold Chemical compound [Au] PCHJSUWPFVWCPO-UHFFFAOYSA-N 0.000 abstract description 11
- 229910052737 gold Inorganic materials 0.000 abstract description 11
- 229910052709 silver Inorganic materials 0.000 abstract description 11
- 239000004332 silver Substances 0.000 abstract description 11
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 abstract description 10
- 229910052751 metal Inorganic materials 0.000 abstract description 6
- 239000002184 metal Substances 0.000 abstract description 6
- 238000003723 Smelting Methods 0.000 abstract description 4
- 238000011161 development Methods 0.000 abstract description 4
- 238000009854 hydrometallurgy Methods 0.000 abstract description 3
- 238000002161 passivation Methods 0.000 abstract description 3
- 238000009279 wet oxidation reaction Methods 0.000 abstract description 3
- 150000002739 metals Chemical class 0.000 abstract description 2
- 239000002244 precipitate Substances 0.000 abstract description 2
- 238000012545 processing Methods 0.000 abstract description 2
- 230000001360 synchronised effect Effects 0.000 abstract description 2
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 abstract 2
- 229910052500 inorganic mineral Inorganic materials 0.000 abstract 2
- 239000011707 mineral Substances 0.000 abstract 2
- 238000011112 process operation Methods 0.000 abstract 1
- 229910052799 carbon Inorganic materials 0.000 description 7
- 238000011084 recovery Methods 0.000 description 7
- 238000001914 filtration Methods 0.000 description 5
- 229910000476 molybdenum oxide Inorganic materials 0.000 description 5
- PQQKPALAQIIWST-UHFFFAOYSA-N oxomolybdenum Chemical compound [Mo]=O PQQKPALAQIIWST-UHFFFAOYSA-N 0.000 description 5
- 229910052569 sulfide mineral Inorganic materials 0.000 description 5
- 230000018109 developmental process Effects 0.000 description 3
- 239000000428 dust Substances 0.000 description 3
- 238000005516 engineering process Methods 0.000 description 3
- 238000004519 manufacturing process Methods 0.000 description 3
- 230000008901 benefit Effects 0.000 description 2
- 230000007547 defect Effects 0.000 description 2
- 239000003814 drug Substances 0.000 description 2
- PNDPGZBMCMUPRI-UHFFFAOYSA-N iodine Chemical compound II PNDPGZBMCMUPRI-UHFFFAOYSA-N 0.000 description 2
- 238000004064 recycling Methods 0.000 description 2
- 241000589902 Leptospira Species 0.000 description 1
- GRYLNZFGIOXLOG-UHFFFAOYSA-N Nitric acid Chemical compound O[N+]([O-])=O GRYLNZFGIOXLOG-UHFFFAOYSA-N 0.000 description 1
- 239000005708 Sodium hypochlorite Substances 0.000 description 1
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 239000003054 catalyst Substances 0.000 description 1
- 238000002485 combustion reaction Methods 0.000 description 1
- 125000004122 cyclic group Chemical group 0.000 description 1
- 238000000280 densification Methods 0.000 description 1
- 238000006056 electrooxidation reaction Methods 0.000 description 1
- 238000010828 elution Methods 0.000 description 1
- 238000005265 energy consumption Methods 0.000 description 1
- 230000007613 environmental effect Effects 0.000 description 1
- 238000003912 environmental pollution Methods 0.000 description 1
- 230000036541 health Effects 0.000 description 1
- 230000006872 improvement Effects 0.000 description 1
- JEIPFZHSYJVQDO-UHFFFAOYSA-N iron(III) oxide Inorganic materials O=[Fe]O[Fe]=O JEIPFZHSYJVQDO-UHFFFAOYSA-N 0.000 description 1
- 239000012633 leachable Substances 0.000 description 1
- 239000000463 material Substances 0.000 description 1
- 230000000813 microbial effect Effects 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 230000003472 neutralizing effect Effects 0.000 description 1
- 229910017604 nitric acid Inorganic materials 0.000 description 1
- 229910052760 oxygen Inorganic materials 0.000 description 1
- 239000001301 oxygen Substances 0.000 description 1
- 239000010970 precious metal Substances 0.000 description 1
- 230000001737 promoting effect Effects 0.000 description 1
- 230000008929 regeneration Effects 0.000 description 1
- 238000011069 regeneration method Methods 0.000 description 1
- 238000011160 research Methods 0.000 description 1
- SUKJFIGYRHOWBL-UHFFFAOYSA-N sodium hypochlorite Chemical compound [Na+].Cl[O-] SUKJFIGYRHOWBL-UHFFFAOYSA-N 0.000 description 1
Images
Classifications
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B23/00—Obtaining nickel or cobalt
- C22B23/04—Obtaining nickel or cobalt by wet processes
- C22B23/0407—Leaching processes
- C22B23/0415—Leaching processes with acids or salt solutions except ammonium salts solutions
- C22B23/043—Sulfurated acids or salts thereof
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
-
- 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
- C22B23/00—Obtaining nickel or cobalt
- C22B23/04—Obtaining nickel or cobalt by wet processes
- C22B23/0407—Leaching processes
- C22B23/0446—Leaching processes with an ammoniacal liquor or with a hydroxide of an alkali or alkaline-earth metal
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2203/00—Specified materials treated by the flotation agents; Specified applications
- B03D2203/02—Ores
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
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- Engineering & Computer Science (AREA)
- Chemical & Material Sciences (AREA)
- Manufacturing & Machinery (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Metallurgy (AREA)
- Organic Chemistry (AREA)
- Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Geology (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention discloses a process method for treating molybdenite by a full wet method under normal pressure, which utilizes the chemical catalytic property of iodine to eliminate the problem of dense passivation film of molybdenum-containing sulphide minerals such as molybdenite in the conventional acid leaching process, promotes the molybdenum-containing sulphide minerals such as molybdenite to react to generate hydrated molybdenum trioxide and elemental sulfur, and precipitates the hydrated molybdenum trioxide into slag phase. In the reduction of SO 2 And the high-efficiency oxidation of molybdenum concentrate under wet conditions is realized while the pollution of flue gas emission and the synchronous separation of copper and molybdenum are realized. Elemental sulfur and noble metals (gold and silver) in the acid leaching slag are further enriched in sulfur concentrate through a flotation process and are used for sale or smelting processing; and the hydrated molybdenum trioxide is enriched in tailings and sent to an ammonia leaching process. The wet oxidation pretreatment process is subsequently connected with the conventional ammonia leaching process operation, so that the full-flow wet treatment process of molybdenite is realized, and simultaneously, the associated valuable metals such as copper, rhenium, gold, silver and the like in the molybdenite are effectively recovered, and the wet oxidation pretreatment process has remarkable significance for the development of the hydrometallurgy of molybdenum.
Description
Technical Field
The invention relates to the field of nonferrous metal hydrometallurgy, in particular to a process method for treating molybdenite by a full wet method under normal pressure.
Background
Along with the increasingly strict requirements of environmental protection and the trend of lean, mixed and refined molybdenum-containing ore, the process has increasingly remarkable defects, mainly comprising the following steps:
(1)SO 2 is an environmental pollution: a large amount of flue gas is generated in the roasting process, the environment is seriously polluted, and the flue gas contains a large amount of metal dust, so that the production environment is seriously polluted, and the health of workers is damaged. At present, although the production process has the procedures of purifying and recovering flue gas, SO generated by roasting 2 And metal-containing dust pollution are not completely eradicated.
(2) Metal comprehensive recovery is poor: in the roasting process, about 3% of molybdenum is lost from flue gas in the form of dust, the associated rare element rhenium is distributed and dispersed in the roasting process, the recovery process is complicated, less manufacturers recover, the recovery rate is generally lower, and only about 70%.
(3) Unsuitable for treating low-grade complex ores: the traditional process is suitable for treating high-grade molybdenum concentrate, has low adaptability to low-grade complex ores, and along with the development of industrial technology, high-grade high-quality molybdenum ore resources are gradually in shortage, the treatment pressure of the traditional process is increased, and the development of the process for efficiently utilizing the low-grade complex molybdenum ore resources is urgent.
The molybdenum concentrate wet treatment process has been studied for 50 years, but has shortcomings. The oxygen autoclaving method and the nitric acid oxidation method need high-temperature and high-pressure equipment, have high equipment requirements, are difficult to control the process conditions of the leaching process, and have certain potential safety hazards. The sodium hypochlorite method has larger medicament consumption, the electro-oxidation method has high cost and energy consumption but low leaching efficiency, and the biological oxidation method is still in the research stage at present.
Disclosure of Invention
Aiming at the defects of the prior art, the invention aims to provide a process method for treating molybdenite by a full-wet method under normal pressure, which takes a non-roasting process as a pre-oxidation means and comprehensively recovers and treats subsequent leaching liquid and slag, thereby realizing the full-flow wet leaching of sulfide minerals such as the molybdenite and reducing SO (sulfur dioxide) 2 And (3) the flue gas is discharged, copper and molybdenum are efficiently separated, and the elemental sulfur and noble metal are enriched and recovered.
In order to achieve the above purpose, the present invention adopts the following technical scheme:
a process method for treating molybdenite by a full wet method under normal pressure condition comprises the following specific processes:
s1, catalytic oxidation acid leaching: mixing molybdenite with sulfuric acid-ferric sulfate mixed solution, adding iodide ions, and stirring at 70-95 ℃ for reaction to enable the molybdenite to react to generate hydrated molybdenum trioxide and elemental sulfur;
s2, solid-liquid separation of the oxidation acid leaching ore pulp: carrying out solid-liquid separation on the oxidized acid leaching ore pulp obtained after the reaction in the step S1 to obtain acid leaching liquid and acid leaching slag, and washing the acid leaching slag; step S3, acid leaching liquid enters a step S4, and acid leaching slag enters a step S;
s3, adsorption of iodine: blowing air into the pickle liquor obtained in the step S2, converting free iodine ions in the pickle liquor into iodine simple substances, and adsorbing the iodine simple substances on an activated carbon column when the pickle liquor flows through the activated carbon column to finally obtain acidic iodine-removed liquor with iodine content lower than 1 ppm; the acidic iodine-removed solution enters a step S9;
s4, acid leaching residue floatation: carrying out flotation on the acid leaching slag washed in the step S2 to obtain sulfur concentrate and tailings, wherein elemental sulfur and noble metals are enriched in the sulfur concentrate, and hydrated molybdenum trioxide enters the tailings;
s5, ammonia leaching of flotation tailings: mixing the tailings obtained by the flotation in the step S4 with ammonia water, stirring and reacting at 25-40 ℃ to generate soluble ammonium molybdate;
s6, removing impurities from ammonia leaching liquid: stirring the ammonia leaching pulp obtained in the step S5 at a constant temperature of 35-40 ℃, slowly adding the prepared ammonium sulfide solution into the ammonia leaching pulp for 0.5-2h until the supernatant is colorless and transparent;
s7, solid-liquid separation of impurity-removed ore pulp: performing solid-liquid separation on the impurity-removed ore pulp obtained in the step S6 to obtain ammonia leaching solution and ammonia leaching residue, and washing the ammonia leaching residue;
s8, ammonium molybdate preparation: evaporating and crystallizing the ammonia immersion liquid obtained in the step S7 to obtain an ammonium molybdate product with the purity reaching the standard;
s9, recovering rhenium: the acidic iodine-removing liquid obtained in the step S3 is subjected to extraction, back extraction, evaporation and crystallization to obtain an ammonium rhenate product;
s10, recovering copper: extracting, back-extracting and electrodepositing the rhenium extraction raffinate obtained in the step S9 to obtain cathode copper;
s11, biological oxidation: and (3) delivering the copper extraction raffinate obtained in the step (S10) to a reaction tank rich in thiobacillus ferrooxidans, blowing air, and oxidizing ferrous ions in the copper extraction raffinate into ferric ions under the action of microorganisms to obtain an acidic ferric sulfate solution.
Further, in step S1, the amount of iodine ions added is 100ppm to 500ppm.
Further, in the step S1, the liquid-solid ratio of the molybdenite and the mixed solution is 3:1-5:1, the stirring reaction time is 6-8 hours, the pH value of the solution is controlled to be 1-1.8, and the oxidation-reduction potential is 820mV-875mV vs.
Further, in the step S3, the pH of the pickle liquor is controlled to be 1-1.8.
Further, in the step S3, the speed of the pickling liquid flowing through the activated carbon column is controlled to be 0.2-5L/S.
Further, in the step S4, most of sulfur concentrate obtained by flotation is sold, and a small part of sulfur concentrate is combusted to prepare sulfur dioxide gas, combustion flue gas is washed by sodium hydroxide to obtain sodium sulfite solution, and the sodium sulfite solution is adopted to desorb the elemental iodine in the iodine-carrying activated carbon column obtained in the step S3 into iodide ions.
Further, in the step S4, the liquid-solid ratio mL/g of the ammonia leaching reaction is 3:1-5:1, and the stirring reaction time is 0.5-2h.
Further, in step S11, in the biological oxidation, the pH is controlled to be 1-1.8, and the oxidation-reduction potential is 800mV-880mV vs.
Further, the iodine ion-containing solution obtained by desorption and the acidic ferric sulfate solution obtained in the step S11 are mixed and returned to the step S1 for use.
Further, carrying out open-circuit treatment on a part of the acidic ferric sulfate solution obtained in the step S11, neutralizing the acidic ferric sulfate solution by limestone powder, and controlling the reaction pH to 7-8; after solid-liquid separation, the neutralized slag is transported to a neutralized slag warehouse for storage, and the neutralized liquid is recycled or discharged.
The invention has the beneficial effects that:according to the invention, by utilizing the chemical catalytic property of iodine, trace iodide ions are added into the sulfuric acid-ferric sulfate mixed solution under normal pressure, so that the problem of passivation film densification of molybdenum-containing sulfide minerals such as molybdenite in the conventional acid leaching process can be solved, the molybdenum-containing sulfide minerals such as molybdenite can be promoted to react to generate hydrated molybdenum trioxide and elemental sulfur, and the hydrated molybdenum trioxide is precipitated into a slag phase. The process reduces SO 2 The high-efficiency oxidation of molybdenum concentrate under wet conditions can be realized while the pollution of flue gas emission and the synchronous separation of copper and molybdenum are realized. Then, elemental sulfur and noble metals (gold and silver) in the acid leaching slag are further enriched in sulfur concentrate through a floatation process for sale or smelting processing; and the hydrated molybdenum trioxide is enriched in tailings and sent to an ammonia leaching process. The subsequent ammonia leaching operation of the wet oxidation process realizes the full-flow wet treatment process of molybdenite, and simultaneously effectively recovers the associated valuable metals such as copper, rhenium, gold, silver and the like in the molybdenite, thereby having remarkable significance for the development of the hydrometallurgy of molybdenum.
In addition, in the invention, the cyclic use of iodide ions can be realized by utilizing an activated carbon adsorption-sodium sulfite elution process; ferric ions in the copper extraction raffinate are subjected to the action of microorganisms such as thiobacillus ferrooxidans, so that low-cost sustainable environment-friendly ferric ion regeneration is realized; the flotation technology is utilized to realize the separation of molybdenum, gold, silver and other noble metals, the elemental sulfur, gold, silver and other noble metals are enriched in sulfur concentrate, the sulfur concentrate can be directly sold or smelted and processed, the molybdenum grade in tailings is further improved, and the subsequent molybdenum extraction and product preparation are facilitated; the flotation sulfur concentrate can be roasted to prepare sulfur dioxide for iodine desorption; the leached iodine ions are adsorbed on the activated carbon and desorbed by sulfur dioxide, and are used for subsequent leaching, so that the recycling and reuse of iodine are realized; meanwhile, the gold and silver grade of the calcine is continuously improved, and the calcine can be directly sold or subjected to gold and silver recovery.
The invention relates to the recycling of key materials such as iodine, ferric ions, active carbon, sulfur dioxide, extractant and the like, and has the advantages of low medicament consumption and low production cost, and the application of the whole process technology in the field of comprehensive recovery of molybdenite is the first time, thereby having important significance for promoting the improvement of the whole technical level in the molybdenum smelting industry.
Drawings
FIG. 1 is a schematic flow chart of the method of examples 1-5 of the present invention.
Detailed Description
The present invention will be further described with reference to the accompanying drawings, and it should be noted that, while the present embodiment provides a detailed implementation and a specific operation process on the premise of the present technical solution, the protection scope of the present invention is not limited to the present embodiment.
Example 1
As shown in FIG. 1, 1kg of molybdenite containing Mo 39.01%, cu 4.20%, S32.75%, au 8.6g/t and Ag 114g/t (molybdenum sulfide is 100% and molybdenum trioxide is 0%) was placed in a closed reactor, a sulfuric acid-ferric sulfate mixed solution (pH=1.2) was added at a ratio of liquid-solid ratio mL/g of 5:1, the solution potential was 875mV (vs. SHE), and the iodine content of the solution was 500ppm (I) - Ion meter), stirring and reacting for 8 hours at the normal pressure and 95 ℃ to enable molybdenite to react to generate hydrated molybdenum trioxide and elemental sulfur element, and copper, rhenium and other leachable impurity elements to be dissolved in the pickle liquor.
After washing by liquid-solid separation, the residue was dried at 60℃for 8 hours in a dry box to obtain 924.51g (Mo 42.20%, cu 0.23%, S35.07%) of acid leaching residue. Wherein the molybdenum sulfide accounts for 0.65 percent, the molybdenum oxide accounts for 99.35 percent, and the molybdenum oxidation rate is 99.35 percent. The pickle liquor is oxidized by air and flows through an active carbon column, and the finally obtained acid iodine-removed liquor contains 1ppm of iodine.
Selecting 900g of acid leaching residues, washing, and then sending to a flotation process to obtain 337.03g (Mo 1.82%, cu 0.58%, S88.97%, au 23.6g/t and Ag 313 g/t) of sulfur concentrate, wherein the yield is 37.45%; 562.97g (66.37% Mo, 0.02% Cu, 2.80% S, 0.7g/t Au and 10g/t Ag) of tailings are obtained and sent to the subsequent ammonia leaching process.
550g of flotation tailings can reach 96.26 percent of Mo leaching rate under the conditions of leaching temperature of 25 ℃, ammonia water dosage of 1.3 times of theoretical quantity, liquid-solid ratio of 5:1 (mL: g) and reaction time of 60 min. Stirring the ammonia leaching pulp at a constant temperature of 35 ℃, and slowly adding the prepared ammonium sulfide solution (10 g/L) into the ammonia leaching pulp for 1h. After the reaction is finished, separating slag and liquid, and filtering to obtain 198.52g (Mo 6.88%, cu 0.002%, S7.77%, au 2g/t and Ag 27 g/t) of ammonia leaching slag, wherein the slag rate is 36.10%; the ammonia immersion liquid is evaporated and crystallized to obtain an ammonium molybdate product, and the purity of the ammonium molybdate product can reach 99.97 percent.
And (3) extracting rhenium from the acidic iodine-removed solution by adopting a rhenium extractant (the extraction rate is more than 99 percent, and the rhenium concentration in raffinate is lower than 0.05 ppm), washing for three times, back-extracting by adopting ammonia water, and evaporating and crystallizing to obtain crude ammonium rhenate (Re 64.72 percent). And (3) carrying out two-stage copper extraction on the rhenium raffinate by adopting a copper extractant (the extraction rate is more than 95%), washing for one time, carrying out one-stage back extraction on the rhenium raffinate by adopting an electrodeposited lean solution, and then conveying the rhenium raffinate to an electrodeposited workshop to obtain a cathode copper plate (more than 99.95%). Finally, the copper extraction raffinate flows into a microbial oxidation pond, and ferrous ions are oxidized into ferric ions under the action of microorganisms such as pH 1-1.8, thiobacillus ferrooxidans, leptospira ferrooxidans and the like, and the solution potential is raised to be 750 mv vs.
And (3) introducing 0.5mol/L NaOH solution into sulfur dioxide gas generated by burning part of sulfur concentrate to obtain sodium sulfite solution. The sodium sulfite solution circularly washes the iodine-carrying active carbon column, reduces the iodine simple substance adsorbed on the active carbon into iodine ions, and continuously enriches the iodine ions. The iodine ion solution is acidified to adjust the pH value to about 1.5, and the iodine ion solution and the regenerated ferric iron solution are mixed and returned to the leaching process.
Example 2
As shown in FIG. 1, 1kg of molybdenite containing Mo 39.01%, cu 4.20%, S32.75%, au 8.6g/t and Ag 114g/t (molybdenum sulfide is 100% and molybdenum trioxide is 0%) was placed in a closed reactor, a sulfuric acid-ferric sulfate mixed solution (pH=1.2) was added at a liquid-solid ratio of mL/g 5:1, the solution potential was 875mV (vs. SHE), and the iodine content of the solution was 300ppm (in I) - Ion meter), stirring and reacting for 8 hours at the normal pressure and 90 ℃;
after liquid-solid separation and washing, the residue was dried at 60℃for 8 hours in a dry box to give 906.90g (Mo 43.01%, cu 0.32%, S33.58%, au 9.5g/t, ag 125 g/t) of acid leaching residue. Wherein the molybdenum sulfide accounts for 0.97 percent, the molybdenum oxide accounts for 99.03 percent, and the molybdenum oxidation rate is 99.03 percent. The pickle liquor was subjected to air oxidation and then passed through an activated carbon column with an iodine content of 1ppm.
Selecting 900g of acid leaching residues, washing, and then sending to a flotation process to obtain 327.71g (Mo 2.26%, cu 0.85%, S87.62%, au 24.7g/t and Ag 327 g/t) of sulfur concentrate, wherein the yield is 36.41%; 572.29g (66.35% Mo, 0.03% Cu, 2.64% S, 0.7g/t Au and 10g/t Ag) of tailings are obtained and sent to the subsequent ammonia leaching process. 550g of flotation tailings can reach 94.48 percent of Mo leaching rate under the conditions of leaching temperature of 25 ℃, ammonia water dosage of 1.3 times of theoretical quantity, liquid-solid ratio of 5:1 (mL: g) and reaction time of 60 min. Stirring the ammonia leaching pulp at a constant temperature of 40 ℃, and slowly adding the prepared ammonium sulfide solution (10 g/L) into the ammonia leaching pulp for 1h. After the reaction is finished, separating slag and liquid, filtering to obtain 205.07g of ammonia leaching slag (Mo 9.82%, cu 0.004%, S7.08%, au 2.0g/t and Ag26 g/t) with slag rate of 37.29%; the ammonia immersion liquid is evaporated and crystallized to obtain an ammonium molybdate product, and the purity of the ammonium molybdate product can reach 99.96 percent.
Subsequent procedure preparation was as in example 1.
Example 3
As shown in FIG. 1, 1kg of molybdenite containing Mo 39.01%, cu 4.20%, S32.75%, au 8.6g/t and Ag 114g/t (molybdenum sulfide is 100% and molybdenum trioxide is 0%) was placed in a closed reactor, a sulfuric acid-ferric sulfate mixed solution (pH=1.2) was added at a liquid-solid ratio of mL/g 4:1, the solution potential was 850mV (vs. SHE), and the iodine content of the solution was 400ppm (in I) - Ion meter), stirring and reacting for 7h at the normal pressure and 85 ℃; after liquid-solid separation and washing, the residue was dried at 60℃for 8 hours in a dry box to give 915.27g (Mo 42.62%, cu 0.41%, S33.99%, au 9.4g/t, ag 124 g/t) of acid leaching residue. Wherein the molybdenum sulfide accounts for 1.13 percent, the molybdenum oxide accounts for 98.87 percent, and the molybdenum oxidation rate is 98.87 percent. The pickle liquor is oxidized by air, flows through an active carbon column, and the iodine content of the pickle liquor after the acid iodine removal is 1ppm.
Selecting 900g of acid leaching residues, washing, and then sending to a flotation process, wherein 332.89g (Mo 2.38%, cu 1.06%, S87.31%, au 24g/t and Ag 320 g/t) of sulfur concentrate is selected, and the yield is 36.99%; 567.11g (Mo 66.25%, cu 0.03%, S2.70%, au 0.7g/t, ag 10 g/t) of tailings are obtained and sent to the subsequent ammonia leaching process. 550g of flotation tailings can reach 95.77 percent of Mo leaching rate under the conditions of leaching temperature of 25 ℃, ammonia water dosage of 1.3 times of theoretical quantity, liquid-solid ratio of 5:1 (mL: g) and reaction time of 60 min. Stirring the ammonia leaching pulp at a constant temperature of 40 ℃, and slowly adding the prepared ammonium sulfide solution (10 g/L) into the ammonia leaching pulp for 1h. After the reaction is finished, separating slag and liquid, and filtering to obtain 200.89g of ammonia leaching slag (Mo 7.67%, cu 0.004%, S7.38%, au 2g/t and Ag 27 g/t) with a slag rate of 35.53%; the ammonia immersion liquid is evaporated and crystallized to obtain an ammonium molybdate product, and the purity of the ammonium molybdate product can reach 99.95 percent.
Subsequent procedure preparation was as in example 1.
Example 4
As shown in FIG. 1, 1kg of molybdenite containing Mo 39.01%, cu 4.20%, S32.75%, au 8.6g/t and Ag 114g/t (molybdenum sulfide is 100% and molybdenum trioxide is 0%) was placed in a closed reactor, a sulfuric acid-ferric sulfate mixed solution (pH=1.2) was added at a liquid-solid ratio of mL/g 3:1, the solution potential was 850mV (vs. SHE), and the iodine content of the solution was 400ppm (in I) - Ion meter), stirring and reacting for 6h at the normal pressure and 85 ℃; after liquid-solid separation and washing, the slag was dried at 60℃for 8 hours in a dry box to obtain 920.17g (Mo 42.39%, cu 0.45%, S34.17%, au 9.3g/t, ag 124 g/t) of acid leaching slag. Wherein the molybdenum sulfide accounts for 1.35 percent, the molybdenum oxide accounts for 98.64 percent, and the molybdenum oxidation rate is 98.64 percent. The pickle liquor is oxidized by air, flows through an active carbon column, and the iodine content of the pickle liquor after the acid iodine removal is 1ppm.
Selecting 900g of acid leaching residues, washing, and then sending to a flotation process to obtain 335.07g (Mo 2.59%, cu 1.16%, S86.93%, au 23.8g/t and Ag 315 g/t) of sulfur concentrate, wherein the yield is 37.34%; 563.93g (66.12% Mo, 0.04% Cu, 2.73% S, 0.7g/t Au and 10g/t Ag) of tailings are obtained and sent to the subsequent ammonia leaching process. 550g of flotation tailings can reach 95.84 percent of Mo leaching rate under the conditions of leaching temperature of 25 ℃, ammonia water dosage of 1.3 times of theoretical quantity, liquid-solid ratio of 5:1 (mL: g) and reaction time of 60 min. Stirring the ammonia leaching pulp at a constant temperature of 40 ℃, and slowly adding the prepared ammonium sulfide solution (10 g/L) into the ammonia leaching pulp for 1h. After the reaction is finished, separating slag and liquid, and filtering to obtain 201.29g of ammonia leaching slag (Mo 7.52%, cu 0.004%, S7.45%, au 2g/t and Ag 27 g/t) with a slag rate of 35.50%; the ammonia immersion liquid is evaporated and crystallized to obtain an ammonium molybdate product, and the purity of the ammonium molybdate product can reach 99.94 percent.
Subsequent procedure preparation was as in example 1.
Example 5
As shown in FIG. 1, 1kg of molybdenite containing Mo 39.01%, cu 4.20%, S32.75%, au 8.6g/t and Ag 114g/t (molybdenum sulfide is 100% and molybdenum trioxide is 0%) was placed in a closed reactor, a sulfuric acid-ferric sulfate mixed solution (pH=1.2) was added at a liquid-solid ratio of mL/g 3:1, the solution potential was 820mV (vs. SHE), and the iodine content of the solution was 300ppm (in I) - Ion meter), stirring and reacting for 8 hours at the normal pressure and 70 ℃; after liquid-solid separation and washing, the slag was dried at 60℃for 8 hours in a dry box to obtain 928.76g (Mo 42.00%, cu 0.54%, S34.56%, au 9.2g/t, ag 123 g/t) of acid leaching slag. Wherein the molybdenum sulfide accounts for 1.78 percent, the molybdenum oxide accounts for 98.22 percent, and the molybdenum oxidation rate is 98.22 percent. The pickle liquor is oxidized by air, flows through an active carbon column, and the iodine content of the pickle liquor after the acid iodine removal is 1ppm.
Selecting 900g of acid leaching residues, washing, and then sending to a flotation process to obtain 341.94g (Mo 2.95%, cu 1.36%, S86.41%, au 23.1g/t and Ag 307 g/t) of sulfur concentrate, wherein the yield is 37.99%; 558.06g (Mo 65.93%, cu 0.04%, S2.79%, au 0.7g/t, ag 10 g/t) of tailings are obtained and sent to the subsequent ammonia leaching process. 550g of flotation tailings can reach 94.48 percent of Mo leaching rate under the conditions of leaching temperature of 40 ℃, ammonia water dosage of 1.3 times theoretical amount, liquid-solid ratio of 3:1 (mL: g) and reaction time of 60 min. Stirring the ammonia leaching pulp at a constant temperature of 40 ℃, and slowly adding the prepared ammonium sulfide solution (10 g/L) into the ammonia leaching pulp for 1h. After the reaction is finished, separating slag and liquid, and filtering to obtain 200.33g of ammonia leaching slag (Mo 6.57%, cu 0.004%, S7.55%, au 2g/t and Ag 27 g/t) with a slag rate of 36.42%; the ammonia immersion liquid is evaporated and crystallized to obtain an ammonium molybdate product, and the purity of the ammonium molybdate product can reach 99.93 percent.
Subsequent procedure preparation was as in example 1.
TABLE 1 molybdenum concentrate treatment results for certain molybdenum smelting enterprises
Based on the catalytic property of iodine on chemical reaction, the above examples 1-5 can eliminate the problem of dense passivation film of molybdenum-containing sulfide minerals such as molybdenite in the conventional acid leaching process by adding trace iodide ions into the sulfuric acid-ferric sulfate mixed solution under normal pressure, promote the molybdenum-containing sulfide minerals such as molybdenite to react to generate hydrated molybdenum trioxide and elemental sulfur,
MoS 2 +6Fe 3+ +4H 2 O→MoO 3 ·H 2 O+2S+6Fe 2+ +6H +
and (3) the hydrated molybdenum trioxide precipitate enters a slag phase, is subjected to dense washing, and enters a conventional flotation process to obtain sulfur concentrate and flotation tailings. Wherein, precious metals such as elemental sulfur, gold and silver enter sulfur concentrate through floatation, and further enrichment and recovery are obtained. The sulfur concentrate can be used for selling or roasting, the generated sulfur dioxide is used for desorbing the iodine catalyst, and meanwhile, the gold and silver grade of the roasted product is continuously improved, and the sulfur concentrate can be directly sold or subjected to gold and silver recovery. And the flotation tailings enter an ammonia leaching process and react with ammonia water to generate ammonium molybdate, and rare noble metals in ammonia leaching residues are continuously enriched. Copper and rhenium are separated from the molybdenum by entering the solution during the catalytic oxidation acid leaching process. In summary, the process of examples 1-5 has significant economic and social benefits.
Various modifications and variations of the present invention will be apparent to those skilled in the art in light of the foregoing teachings and are intended to be included within the scope of the following claims.
Claims (8)
1. A process method for treating molybdenite by a full wet method under normal pressure is characterized by comprising the following specific steps:
s1, catalytic oxidation acid leaching: mixing molybdenite with sulfuric acid-ferric sulfate mixed solution, adding iodide ions, and stirring at 70-95 ℃ for reaction to enable the molybdenite to react to generate hydrated molybdenum trioxide and elemental sulfur; the liquid-solid ratio mL/g of the molybdenite and sulfuric acid-ferric sulfate mixed solution is 3:1-5:1, the stirring reaction time is 6-8h, the pH value of the solution is controlled to be 1-1.8, and the oxidation-reduction potential is 820mV-875mV vs. SHE;
s2, solid-liquid separation of the oxidation acid leaching ore pulp: carrying out solid-liquid separation on the oxidized acid leaching ore pulp obtained after the reaction in the step S1 to obtain acid leaching liquid and acid leaching slag, and washing the acid leaching slag; step S3, acid leaching liquid enters a step S4, and acid leaching slag enters a step S;
s3, adsorption of iodine: blowing air into the pickle liquor obtained in the step S2, converting free iodine ions in the pickle liquor into iodine simple substances, and adsorbing the iodine simple substances on an activated carbon column when the pickle liquor flows through the activated carbon column to finally obtain acidic iodine-removed liquor with iodine content lower than 1 ppm; the acidic iodine-removed solution enters a step S9;
s4, acid leaching residue floatation: carrying out flotation on the acid leaching slag washed in the step S2 to obtain sulfur concentrate and tailings, wherein elemental sulfur and noble metals are enriched in the sulfur concentrate, and hydrated molybdenum trioxide enters the tailings;
s5, ammonia leaching of flotation tailings: mixing the tailings obtained by the flotation in the step S4 with ammonia water, stirring and reacting at 25-40 ℃ to generate soluble ammonium molybdate;
s6, removing impurities from ammonia leaching liquid: stirring the ammonia leaching pulp obtained in the step S5 at a constant temperature of 35-40 ℃, slowly adding the prepared ammonium sulfide solution into the ammonia leaching pulp for 0.5-2h until the supernatant is colorless and transparent;
s7, solid-liquid separation of impurity-removed ore pulp: performing solid-liquid separation on the impurity-removed ore pulp obtained in the step S6 to obtain ammonia leaching solution and ammonia leaching residue, and washing the ammonia leaching residue;
s8, ammonium molybdate preparation: evaporating and crystallizing the ammonia immersion liquid obtained in the step S7 to obtain an ammonium molybdate product with the purity reaching the standard;
s9, recovering rhenium: the acidic iodine-removing liquid obtained in the step S3 is subjected to extraction, back extraction, evaporation and crystallization to obtain an ammonium rhenate product;
s10, recovering copper: extracting, back-extracting and electrodepositing the rhenium extraction raffinate obtained in the step S9 to obtain cathode copper;
s11, biological oxidation: delivering the copper extraction raffinate obtained in the step S10 to a reaction tank rich in thiobacillus ferrooxidans, blowing air, and oxidizing ferrous ions in the copper extraction raffinate into ferric ions under the action of microorganisms to obtain an acidic ferric sulfate solution; the pH is controlled to be 1-1.8, and the oxidation-reduction potential is 800mV-880mV vs.
2. The process according to claim 1, wherein in step S1, the amount of iodine ions added is 100ppm to 500ppm.
3. The process according to claim 1, wherein in step S3, the pH of the pickling solution is controlled to be 1-1.8.
4. The process according to claim 1, wherein in step S3, the velocity of the pickling liquid flowing through the activated carbon column is controlled to be 0.2-5L/S.
5. The process according to claim 1, wherein in step S4, a major part of the sulfur concentrate obtained by flotation is sold and a minor part is burned to produce sulfur dioxide gas, the burned flue gas is washed with sodium hydroxide to obtain sodium sulfite solution, and the sodium sulfite solution is used to desorb elemental iodine in the iodine-carrying activated carbon column obtained in step S3 into iodide ions.
6. The process according to claim 1, wherein in step S5, the liquid-solid ratio mL/g of the ammonia leaching reaction is 3:1-5:1, and the stirring reaction time is 0.5-2h.
7. The process according to claim 5, wherein the iodine ion-containing solution obtained by desorption and the acidic ferric sulfate solution obtained in step S11 are mixed and returned to step S1 for use.
8. The process according to claim 1, wherein a part of the acidic ferric sulfate solution obtained in step S11 is subjected to an open-circuit treatment, the acidic ferric sulfate solution is neutralized with limestone powder, and the reaction pH is controlled to 7-8; after solid-liquid separation, the neutralized slag is transported to a neutralized slag warehouse for storage, and the neutralized liquid is recycled or discharged.
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