CN118186205A - Method for recovering vanadium and iron from vanadium shale - Google Patents
Method for recovering vanadium and iron from vanadium shale Download PDFInfo
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- CN118186205A CN118186205A CN202410520780.6A CN202410520780A CN118186205A CN 118186205 A CN118186205 A CN 118186205A CN 202410520780 A CN202410520780 A CN 202410520780A CN 118186205 A CN118186205 A CN 118186205A
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- 229910052720 vanadium Inorganic materials 0.000 title claims abstract description 357
- LEONUFNNVUYDNQ-UHFFFAOYSA-N vanadium atom Chemical compound [V] LEONUFNNVUYDNQ-UHFFFAOYSA-N 0.000 title claims abstract description 354
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 title claims abstract description 164
- 238000000034 method Methods 0.000 title claims abstract description 92
- 229910052742 iron Inorganic materials 0.000 title claims abstract description 68
- 239000007788 liquid Substances 0.000 claims abstract description 327
- 238000002386 leaching Methods 0.000 claims abstract description 194
- 239000002253 acid Substances 0.000 claims abstract description 149
- 238000002156 mixing Methods 0.000 claims abstract description 108
- 235000021110 pickles Nutrition 0.000 claims abstract description 77
- 239000000243 solution Substances 0.000 claims abstract description 75
- 230000001105 regulatory effect Effects 0.000 claims abstract description 74
- 239000002893 slag Substances 0.000 claims abstract description 66
- 239000000843 powder Substances 0.000 claims abstract description 63
- 239000003795 chemical substances by application Substances 0.000 claims abstract description 51
- 229910052731 fluorine Inorganic materials 0.000 claims abstract description 51
- YCKRFDGAMUMZLT-UHFFFAOYSA-N Fluorine atom Chemical compound [F] YCKRFDGAMUMZLT-UHFFFAOYSA-N 0.000 claims abstract description 50
- GNTDGMZSJNCJKK-UHFFFAOYSA-N divanadium pentaoxide Chemical compound O=[V](=O)O[V](=O)=O GNTDGMZSJNCJKK-UHFFFAOYSA-N 0.000 claims abstract description 50
- 239000011737 fluorine Substances 0.000 claims abstract description 50
- OAICVXFJPJFONN-UHFFFAOYSA-N Phosphorus Chemical compound [P] OAICVXFJPJFONN-UHFFFAOYSA-N 0.000 claims abstract description 43
- 239000011574 phosphorus Substances 0.000 claims abstract description 43
- 229910052698 phosphorus Inorganic materials 0.000 claims abstract description 43
- 238000005406 washing Methods 0.000 claims abstract description 41
- 238000001914 filtration Methods 0.000 claims abstract description 39
- 239000011550 stock solution Substances 0.000 claims abstract description 38
- BZSXEZOLBIJVQK-UHFFFAOYSA-N 2-methylsulfonylbenzoic acid Chemical compound CS(=O)(=O)C1=CC=CC=C1C(O)=O BZSXEZOLBIJVQK-UHFFFAOYSA-N 0.000 claims abstract description 35
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims abstract description 32
- 229910001868 water Inorganic materials 0.000 claims abstract description 32
- 239000003638 chemical reducing agent Substances 0.000 claims abstract description 29
- 150000003863 ammonium salts Chemical class 0.000 claims abstract description 25
- WBJZTOZJJYAKHQ-UHFFFAOYSA-K iron(3+) phosphate Chemical compound [Fe+3].[O-]P([O-])([O-])=O WBJZTOZJJYAKHQ-UHFFFAOYSA-K 0.000 claims abstract description 19
- 229910000398 iron phosphate Inorganic materials 0.000 claims abstract description 17
- 238000000498 ball milling Methods 0.000 claims abstract description 13
- 230000001376 precipitating effect Effects 0.000 claims abstract description 13
- QAOWNCQODCNURD-UHFFFAOYSA-N sulfuric acid Substances OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 151
- 239000012074 organic phase Substances 0.000 claims description 132
- 238000011084 recovery Methods 0.000 claims description 69
- 238000000605 extraction Methods 0.000 claims description 62
- 238000003756 stirring Methods 0.000 claims description 60
- 239000003011 anion exchange membrane Substances 0.000 claims description 55
- 239000012528 membrane Substances 0.000 claims description 52
- NBIIXXVUZAFLBC-UHFFFAOYSA-N Phosphoric acid Chemical compound OP(O)(O)=O NBIIXXVUZAFLBC-UHFFFAOYSA-N 0.000 claims description 48
- 239000002002 slurry Substances 0.000 claims description 48
- 239000012071 phase Substances 0.000 claims description 45
- 238000005341 cation exchange Methods 0.000 claims description 40
- 238000005273 aeration Methods 0.000 claims description 39
- 238000001556 precipitation Methods 0.000 claims description 30
- 238000005191 phase separation Methods 0.000 claims description 29
- QGZKDVFQNNGYKY-UHFFFAOYSA-O Ammonium Chemical compound [NH4+] QGZKDVFQNNGYKY-UHFFFAOYSA-O 0.000 claims description 25
- 239000000203 mixture Substances 0.000 claims description 25
- 239000012452 mother liquor Substances 0.000 claims description 25
- IJGRMHOSHXDMSA-UHFFFAOYSA-N Atomic nitrogen Chemical compound N#N IJGRMHOSHXDMSA-UHFFFAOYSA-N 0.000 claims description 24
- 229910000147 aluminium phosphate Inorganic materials 0.000 claims description 24
- 239000012492 regenerant Substances 0.000 claims description 24
- 238000004064 recycling Methods 0.000 claims description 22
- PMZURENOXWZQFD-UHFFFAOYSA-L Sodium Sulfate Chemical compound [Na+].[Na+].[O-]S([O-])(=O)=O PMZURENOXWZQFD-UHFFFAOYSA-L 0.000 claims description 16
- PUZPDOWCWNUUKD-UHFFFAOYSA-M sodium fluoride Chemical compound [F-].[Na+] PUZPDOWCWNUUKD-UHFFFAOYSA-M 0.000 claims description 16
- 229910052938 sodium sulfate Inorganic materials 0.000 claims description 16
- 235000011152 sodium sulphate Nutrition 0.000 claims description 16
- 239000003607 modifier Substances 0.000 claims description 15
- NROKBHXJSPEDAR-UHFFFAOYSA-M potassium fluoride Chemical compound [F-].[K+] NROKBHXJSPEDAR-UHFFFAOYSA-M 0.000 claims description 14
- NLXLAEXVIDQMFP-UHFFFAOYSA-N Ammonia chloride Chemical compound [NH4+].[Cl-] NLXLAEXVIDQMFP-UHFFFAOYSA-N 0.000 claims description 12
- ZAMOUSCENKQFHK-UHFFFAOYSA-N Chlorine atom Chemical compound [Cl] ZAMOUSCENKQFHK-UHFFFAOYSA-N 0.000 claims description 12
- 239000000460 chlorine Substances 0.000 claims description 12
- 229910052801 chlorine Inorganic materials 0.000 claims description 12
- 230000009615 deamination Effects 0.000 claims description 12
- 238000006481 deamination reaction Methods 0.000 claims description 12
- 239000007789 gas Substances 0.000 claims description 12
- 239000003350 kerosene Substances 0.000 claims description 12
- 229910052757 nitrogen Inorganic materials 0.000 claims description 12
- 239000002245 particle Substances 0.000 claims description 12
- 230000001172 regenerating effect Effects 0.000 claims description 12
- 230000008929 regeneration Effects 0.000 claims description 12
- 238000011069 regeneration method Methods 0.000 claims description 12
- 239000007787 solid Substances 0.000 claims description 12
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims description 9
- 239000001301 oxygen Substances 0.000 claims description 9
- 229910052760 oxygen Inorganic materials 0.000 claims description 9
- KWSLGOVYXMQPPX-UHFFFAOYSA-N 5-[3-(trifluoromethyl)phenyl]-2h-tetrazole Chemical compound FC(F)(F)C1=CC=CC(C2=NNN=N2)=C1 KWSLGOVYXMQPPX-UHFFFAOYSA-N 0.000 claims description 8
- WUKWITHWXAAZEY-UHFFFAOYSA-L calcium difluoride Chemical compound [F-].[F-].[Ca+2] WUKWITHWXAAZEY-UHFFFAOYSA-L 0.000 claims description 8
- 229910001634 calcium fluoride Inorganic materials 0.000 claims description 8
- SEGLCEQVOFDUPX-UHFFFAOYSA-N di-(2-ethylhexyl)phosphoric acid Chemical group CCCCC(CC)COP(O)(=O)OCC(CC)CCCC SEGLCEQVOFDUPX-UHFFFAOYSA-N 0.000 claims description 8
- SJWFXCIHNDVPSH-UHFFFAOYSA-N octan-2-ol Chemical group CCCCCCC(C)O SJWFXCIHNDVPSH-UHFFFAOYSA-N 0.000 claims description 8
- 239000011775 sodium fluoride Substances 0.000 claims description 8
- 235000013024 sodium fluoride Nutrition 0.000 claims description 8
- 229910001379 sodium hypophosphite Inorganic materials 0.000 claims description 8
- NCPXQVVMIXIKTN-UHFFFAOYSA-N trisodium;phosphite Chemical compound [Na+].[Na+].[Na+].[O-]P([O-])[O-] NCPXQVVMIXIKTN-UHFFFAOYSA-N 0.000 claims description 8
- ACVYVLVWPXVTIT-UHFFFAOYSA-N phosphinic acid Chemical compound O[PH2]=O ACVYVLVWPXVTIT-UHFFFAOYSA-N 0.000 claims description 7
- 239000011698 potassium fluoride Substances 0.000 claims description 7
- 235000003270 potassium fluoride Nutrition 0.000 claims description 7
- ISIJQEHRDSCQIU-UHFFFAOYSA-N tert-butyl 2,7-diazaspiro[4.5]decane-7-carboxylate Chemical compound C1N(C(=O)OC(C)(C)C)CCCC11CNCC1 ISIJQEHRDSCQIU-UHFFFAOYSA-N 0.000 claims description 7
- STCOOQWBFONSKY-UHFFFAOYSA-N tributyl phosphate Chemical compound CCCCOP(=O)(OCCCC)OCCCC STCOOQWBFONSKY-UHFFFAOYSA-N 0.000 claims description 7
- VHUUQVKOLVNVRT-UHFFFAOYSA-N Ammonium hydroxide Chemical compound [NH4+].[OH-] VHUUQVKOLVNVRT-UHFFFAOYSA-N 0.000 claims description 6
- 235000019270 ammonium chloride Nutrition 0.000 claims description 6
- 235000011114 ammonium hydroxide Nutrition 0.000 claims description 6
- BFNBIHQBYMNNAN-UHFFFAOYSA-N ammonium sulfate Chemical compound N.N.OS(O)(=O)=O BFNBIHQBYMNNAN-UHFFFAOYSA-N 0.000 claims description 6
- 229910052921 ammonium sulfate Inorganic materials 0.000 claims description 6
- 235000011130 ammonium sulphate Nutrition 0.000 claims description 6
- LJKDOMVGKKPJBH-UHFFFAOYSA-N 2-ethylhexyl dihydrogen phosphate Chemical compound CCCCC(CC)COP(O)(O)=O LJKDOMVGKKPJBH-UHFFFAOYSA-N 0.000 claims description 3
- 238000006386 neutralization reaction Methods 0.000 abstract description 39
- 239000003814 drug Substances 0.000 abstract description 19
- 239000000284 extract Substances 0.000 abstract description 3
- 239000000047 product Substances 0.000 description 38
- 239000003513 alkali Substances 0.000 description 12
- 239000002244 precipitate Substances 0.000 description 7
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 description 6
- 238000000909 electrodialysis Methods 0.000 description 6
- 230000003472 neutralizing effect Effects 0.000 description 6
- 238000005554 pickling Methods 0.000 description 6
- 238000000746 purification Methods 0.000 description 6
- ZDFBXXSHBTVQMB-UHFFFAOYSA-N 2-ethylhexoxy(2-ethylhexyl)phosphinic acid Chemical group CCCCC(CC)COP(O)(=O)CC(CC)CCCC ZDFBXXSHBTVQMB-UHFFFAOYSA-N 0.000 description 4
- 239000012535 impurity Substances 0.000 description 4
- 239000003014 ion exchange membrane Substances 0.000 description 4
- 238000000926 separation method Methods 0.000 description 4
- 238000001179 sorption measurement Methods 0.000 description 4
- 235000008733 Citrus aurantifolia Nutrition 0.000 description 3
- 235000011941 Tilia x europaea Nutrition 0.000 description 3
- 239000002585 base Substances 0.000 description 3
- 238000005516 engineering process Methods 0.000 description 3
- 239000004571 lime Substances 0.000 description 3
- 238000010979 pH adjustment Methods 0.000 description 3
- 239000002699 waste material Substances 0.000 description 3
- 239000005955 Ferric phosphate Substances 0.000 description 2
- 238000001354 calcination Methods 0.000 description 2
- 238000006243 chemical reaction Methods 0.000 description 2
- 238000002425 crystallisation Methods 0.000 description 2
- 230000008025 crystallization Effects 0.000 description 2
- 238000003795 desorption Methods 0.000 description 2
- 229940032958 ferric phosphate Drugs 0.000 description 2
- 238000001027 hydrothermal synthesis Methods 0.000 description 2
- 229910052500 inorganic mineral Inorganic materials 0.000 description 2
- 229910000399 iron(III) phosphate Inorganic materials 0.000 description 2
- 229910052935 jarosite Inorganic materials 0.000 description 2
- 230000014759 maintenance of location Effects 0.000 description 2
- 239000010445 mica Substances 0.000 description 2
- 229910052618 mica group Inorganic materials 0.000 description 2
- 235000010755 mineral Nutrition 0.000 description 2
- 239000011707 mineral Substances 0.000 description 2
- 239000012466 permeate Substances 0.000 description 2
- NBIIXXVUZAFLBC-UHFFFAOYSA-K phosphate Chemical compound [O-]P([O-])([O-])=O NBIIXXVUZAFLBC-UHFFFAOYSA-K 0.000 description 2
- 230000008092 positive effect Effects 0.000 description 2
- 150000003681 vanadium Chemical class 0.000 description 2
- 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
- VWBLQUSTSLXQON-UHFFFAOYSA-N N.[V+5] Chemical compound N.[V+5] VWBLQUSTSLXQON-UHFFFAOYSA-N 0.000 description 1
- 229910019142 PO4 Inorganic materials 0.000 description 1
- 229910052782 aluminium Inorganic materials 0.000 description 1
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 description 1
- 229910001579 aluminosilicate mineral Inorganic materials 0.000 description 1
- -1 and finally Substances 0.000 description 1
- 239000003245 coal Substances 0.000 description 1
- 230000007547 defect Effects 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000001035 drying Methods 0.000 description 1
- 230000000694 effects Effects 0.000 description 1
- 229910001385 heavy metal Inorganic materials 0.000 description 1
- 229910052595 hematite Inorganic materials 0.000 description 1
- 239000011019 hematite Substances 0.000 description 1
- UQSXHKLRYXJYBZ-UHFFFAOYSA-N iron oxide Inorganic materials [Fe]=O UQSXHKLRYXJYBZ-UHFFFAOYSA-N 0.000 description 1
- PNXOJQQRXBVKEX-UHFFFAOYSA-N iron vanadium Chemical compound [V].[Fe] PNXOJQQRXBVKEX-UHFFFAOYSA-N 0.000 description 1
- 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 description 1
- LIKBJVNGSGBSGK-UHFFFAOYSA-N iron(3+);oxygen(2-) Chemical compound [O-2].[O-2].[O-2].[Fe+3].[Fe+3] LIKBJVNGSGBSGK-UHFFFAOYSA-N 0.000 description 1
- 229910000360 iron(III) sulfate Inorganic materials 0.000 description 1
- 238000005272 metallurgy Methods 0.000 description 1
- 238000005065 mining Methods 0.000 description 1
- 230000003647 oxidation Effects 0.000 description 1
- 238000007254 oxidation reaction Methods 0.000 description 1
- NDLPOXTZKUMGOV-UHFFFAOYSA-N oxo(oxoferriooxy)iron hydrate Chemical compound O.O=[Fe]O[Fe]=O NDLPOXTZKUMGOV-UHFFFAOYSA-N 0.000 description 1
- 239000010452 phosphate Substances 0.000 description 1
- NIFIFKQPDTWWGU-UHFFFAOYSA-N pyrite Chemical compound [Fe+2].[S-][S-] NIFIFKQPDTWWGU-UHFFFAOYSA-N 0.000 description 1
- 229910052683 pyrite Inorganic materials 0.000 description 1
- 239000011028 pyrite Substances 0.000 description 1
- 229920006395 saturated elastomer Polymers 0.000 description 1
- 239000011734 sodium Substances 0.000 description 1
- 229910052708 sodium Inorganic materials 0.000 description 1
- 239000001488 sodium phosphate Substances 0.000 description 1
- 229910000162 sodium phosphate Inorganic materials 0.000 description 1
- 238000000638 solvent extraction Methods 0.000 description 1
- 239000004575 stone Substances 0.000 description 1
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- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention relates to a method for recovering vanadium and iron from vanadium shale. The technical proposal is as follows: crushing and ball milling vanadium shale, and adding sodium chlorate and fluorine-containing leaching aid to obtain leaching powder; mixing the leached powder with a leaching agent, and leaching to obtain a vanadium-containing pickle liquor and leached slag; adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor to obtain a pretreatment liquor; injecting pretreatment liquid into a No.1 liquid inlet, injecting water or raffinate into a No.2 liquid inlet, switching on a power supply and circulating, obtaining the pretreatment liquid after regulating the pH value at the No.1 liquid outlet, and obtaining recovered acid liquid at the No.2 liquid outlet; recovering acid liquor for preparing leaching agent; aerating, filtering and washing the pretreatment liquid after the pH value is regulated to obtain iron-removing slag and extract stock solution; dehydrating the iron-removing slag to obtain an iron phosphate product; extracting the extracting stock solution to obtain raffinate and vanadium-rich solution; and precipitating vanadium from the vanadium-rich solution by ammonium salt, and deaminizing to obtain vanadium pentoxide. The method can effectively recycle vanadium and iron in the vanadium shale, does not generate neutralization slag, has small medicament consumption and no vanadium loss.
Description
Technical Field
The invention belongs to the technical field of vanadium shale. In particular to a method for recovering vanadium and iron from vanadium shale.
Background
The types of minerals in the shale containing vanadium are complex, the grade of vanadium is generally lower, iron-containing minerals such as pyrite, hematite and the like are associated, wherein vanadium is mainly endowed in a mica aluminosilicate mineral lattice, and the structure is stable. The main flow process for extracting vanadium from shale is to obtain vanadium-containing acid leaching solution by a direct acid leaching method or a roasting acid leaching method, and then obtain vanadium products by processes of alkali neutralization, purification and enrichment, ammonium salt vanadium precipitation and the like. A large amount of sulfuric acid is generally consumed in the leaching process and a fluorine-containing leaching aid is added to strengthen the lattice damage to the vanadium-containing mica so as to improve the leaching rate of vanadium. Thus, the vanadium-containing pickle liquor has a lower pH and contains a large amount of impurities Fe and F. In the alkali neutralization process, although part of Fe can be removed in a precipitation form, the problems that the medicament consumption is large, a large amount of neutralization slag can cause V loss in an entrainment adsorption form, redundant acid in the vanadium-containing acid leaching solution can not be utilized, fe in the vanadium-containing acid leaching solution can not be comprehensively recovered, fe in the solution after pH adjustment is still at a higher level and the like exist. During the purification and enrichment process, too high a Fe content can affect the extraction efficiency of the extractant on vanadium and can produce a raffinate containing a large amount of F. In the mainstream process, the raffinate is only recycled for the washing process of leaching residues and neutralization residues, and F in the raffinate cannot be effectively utilized.
Qi Dong et al (Qi Dong, wang Yi. Exploration of iron removal process of vanadium-containing leachate [ J ]. Nonferrous mining metallurgy, 2015, 31 (3): 37-39) in exploration of iron removal process of vanadium-containing leachate, lime neutralization precipitation method is adopted first, pH value is adjusted to 2, part of Fe is removed in advance, and V, fe is separated by reduction-solvent extraction method. The result shows that as the lime dosage increases, the pH value gradually increases, the amount of the neutralized waste residue also gradually increases, and the vanadium loss rate also increases due to the entrainment adsorption effect of the waste residue; when the pH was controlled at 2.0, the vanadium loss was about 6%. In order to reduce the vanadium loss rate, the neutralization slag is washed later, and finally the vanadium loss rate is reduced to about 1 percent. Although the method can achieve the aim of effectively removing Fe, the lime is used for regulating the pH value, so that not only is the residual sulfuric acid in the vanadium-containing pickle liquor wasted, a large amount of neutralization waste residues are generated and vanadium loss is caused, but also the Fe and fluorine in the vanadium-containing pickle liquor cannot be recycled.
The patent technology of purifying raffinate by utilizing shale vanadium extraction neutralization slag (CN 114345290A) comprises the steps of firstly mixing neutralization slag, phosphate and water in proportion, adjusting the pH value to 8-12 by sodium hydroxide, and then placing the mixture in a reaction kettle for hydrothermal reaction to obtain modified neutralization slag; then F and heavy metal elements in the raffinate are adsorbed by utilizing the modified neutralization slag; finally, mixing saturated modified neutralization slag, water, sodium hydroxide and phosphate, and placing the mixture in a reaction kettle for hydrothermal reaction to obtain desorption solution and regenerated modified neutralization slag, wherein the desorption solution can be used as a fluorine-containing leaching aid after crystallization. The process can utilize partial neutralization slag, has strong impurity removal capability, can recycle F in raffinate and can reduce the addition amount of fluorine-containing leaching aid in the subsequent leaching process, but does not solve the problems that residual sulfuric acid in vanadium-containing acid leaching liquid cannot be utilized, a large amount of neutralization slag is generated, vanadium loss is high, iron cannot be recycled and the consumption of medicament is large due to an alkali neutralization method.
According to the 'stone coal pickle liquor vanadium iron extraction comprehensive recovery method' (CN 102127657A) patent technology, after aluminum is removed by crystallization, the pH value is adjusted to 1.0-3.5 by an alkali neutralization method, a mixture containing vanadium jarosite and ammonium jarosite is separated out by oxidation precipitation, the mixture separated out by precipitation is subjected to alkali leaching and filtration to obtain alkali leaching liquor containing vanadium and alkali leaching slag containing iron, and the obtained alkali leaching liquor is added with sulfuric acid to acidify and precipitate vanadium to obtain red vanadium and vanadium precipitation mother liquor; washing to remove sodium after calcining red vanadium, filtering and drying to obtain refined vanadium; adding sulfuric acid into the alkaline leaching residue to dissolve to prepare polymeric ferric sulfate, or calcining the alkaline leaching residue to obtain ferric oxide. The technology can comprehensively recover vanadium and iron, but needs repeated acid-base neutralization to adjust the pH value, has large medicament consumption, can not utilize residual sulfuric acid in the vanadium-containing pickle liquor, and can not recover fluorine.
In summary, the existing method for recovering vanadium and iron has the problems that residual sulfuric acid in the vanadium-containing pickle liquor cannot be recovered, a large amount of neutralization slag is generated, the vanadium loss is high, iron and fluorine are difficult to recover, and the consumption of medicaments is large.
Disclosure of Invention
The invention aims to overcome the defects of the prior art, and aims to provide a method for recovering vanadium and iron from vanadium shale, which is easy to recover vanadium and iron in the vanadium-containing acid leaching solution, can recover residual sulfuric acid and fluorine-containing leaching aid in the vanadium-containing acid leaching solution, has no neutralization slag generation, has high medicament utilization rate, small medicament consumption and no vanadium loss.
In order to achieve the above purpose, the steps of the technical scheme adopted by the invention are as follows:
Step 1, directly leaching vanadium shale
Step 1.1, preparing a leaching agent
And uniformly mixing concentrated sulfuric acid and water or concentrated sulfuric acid, water and recovered acid liquid according to the volume fraction of sulfuric acid of 10-40%, thereby obtaining the leaching agent.
Step 1.2, preparing leached powder
Crushing the vanadium-containing shale, and ball-milling until the particle size is smaller than 74 mu m and 40-70 wt% of the shale is obtained.
The sodium chlorate, the fluorine-containing leaching aid and the raw ore powder are uniformly mixed according to the mass ratio of (0.01-0.05) to (0.01-0.1) to 1.
Step 1.3, direct acid leaching
And mixing the leaching agent and the leaching powder according to the liquid-solid ratio of 0.8-2.5L/kg to obtain leaching slurry.
Stirring the leached slurry for 6-24 hours at the temperature of 80-98 ℃, filtering, and washing to obtain the vanadium-containing pickle liquor and leached slag.
Step2, reduction of the vanadium-containing pickle liquor
Adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor according to the mol ratio of phosphorus to iron in the pretreatment liquor of (1-5) to 1, and stirring for 1-6 h at the temperature of 60-100 ℃ to prepare the pretreatment liquor.
Step3, adjusting the pH of the pretreatment liquid
Step 3.1, device for adjusting pH of pretreatment liquid
The device for adjusting the pH value of the pretreatment liquid comprises: cathode, anode, membrane stack and DC power supply.
The cathode is connected with the negative electrode of the direct current power supply, the anode is connected with the positive electrode of the direct current power supply, and the cathode and the anode are correspondingly arranged on the right side and the left side of the membrane stack.
The membrane stack sequentially comprises a 1 st anion exchange membrane, a 1 st cation exchange membrane, a 2 nd anion exchange membrane, a 2 nd cation exchange membrane, a 3 rd anion exchange membrane, … … th anion exchange membrane, an nth cation exchange membrane and an n+1th anion exchange membrane from the anode to the cathode; and n is a positive integer of 10-1000.
From the anode to cathode direction: the gap between the anode and the 1 st anion exchange membrane forms an anode electrode chamber, the gap between the 1 st anion exchange membrane and the 1 st cation exchange membrane forms a 1-stage regulating chamber, the gap between the 1 st cation exchange membrane and the 2 nd anion exchange membrane forms an n-stage recovery acid chamber, the gap between the 2 nd anion exchange membrane and the 2 nd cation exchange membrane forms a 2-stage regulating chamber, the gap between the 2 nd cation exchange membrane and the 3 rd anion exchange membrane forms an n-1 stage recovery acid chamber, … …, and so on, the gap between the n-1 st anion exchange membrane and the n-1 st cation exchange membrane forms an n-1 stage regulating chamber, the gap between the n-1 st cation exchange membrane and the n anion exchange membrane forms an n-stage recovery acid chamber, the gap between the n cation exchange membrane and the n+1 th anion exchange membrane forms a 1-stage recovery acid chamber, and the gap between the n+1 th anion exchange membrane and the cathode electrode forms a cathode electrode chamber.
The lower part of the 1-level adjusting chamber, the lower part of the 2-level adjusting chamber, the lower part of the 3-level adjusting chamber, … … and the lower part of the n-level adjusting chamber are respectively communicated with the No.1 liquid inlet, and the upper part of the 1-level adjusting chamber, the upper part of the 2-level adjusting chamber, the upper part of the 3-level adjusting chamber, the upper part of the … … and the upper part of the n-level adjusting chamber are respectively communicated with the No.1 liquid outlet.
The lower part of the 1-level recovery acid chamber, the lower part of the 2-level recovery acid chamber, the lower part of the 3-level recovery acid chamber, … … and the lower part of the n-level recovery acid chamber are respectively communicated with the No.2 liquid inlet, and the upper part of the 1-level recovery acid chamber, the upper part of the 2-level recovery acid chamber, the upper part of the 3-level recovery acid chamber, … … and the upper part of the n-level recovery acid chamber are respectively communicated with the No.2 liquid outlet.
And forming a serial loop of the anode electrode chamber, the 1-stage regulating chamber, the n-stage acid recovering chamber, the 2-stage regulating chamber, the n-1-stage acid recovering chamber, … …, the n-1-stage regulating chamber, the 2-stage acid recovering chamber, the n-stage regulating chamber, the 1-stage acid recovering chamber, the cathode electrode chamber and the direct current power supply under the working state to obtain the device for regulating the pH value of the pretreatment liquid.
Step 3.2, adjusting the pH value of the pretreatment liquid
And (3) respectively injecting sodium sulfate solution into an anode electrode chamber and a cathode electrode chamber of the device for adjusting the pH of the pretreatment liquid, injecting the pretreatment liquid prepared in the step (2) from a No.1 liquid inlet of the device for adjusting the pH of the pretreatment liquid, and injecting water or raffinate from a No. 2 liquid inlet of the device for adjusting the pH of the pretreatment liquid.
And switching on a direct current power supply of the device for regulating the pH value of the pretreatment liquid.
The pretreatment liquid injected from the No. 1 liquid inlet flows through a 1-level adjusting chamber, a 2-level adjusting chamber, a 3-level adjusting chamber, … …, an n-1-level adjusting chamber and an n-level adjusting chamber simultaneously, then flows out from a No. 1 liquid outlet and returns to the No. 1 liquid inlet, and the pretreatment liquid is circularly reciprocated until the pH value of the pretreatment liquid after the pH value is adjusted, which flows out from the No. 1 liquid outlet, is 1.3-2.0, and then enters the aeration tank in the step 3.
The water injected from the No. 2 liquid inlet flows through the 1-grade recovery acid chamber, the 2-grade recovery acid chamber, the 3-grade recovery acid chamber, … …, the n-1 grade recovery acid chamber and the n-grade recovery acid chamber at the same time, and then flows out from the No. 2 liquid outlet and returns to the No. 2 liquid inlet, so that the water is circularly reciprocated until the pretreatment liquid flows out from the No. 1 liquid outlet after the pH value is regulated and enters the aeration tank in the step 3, and then the acid liquid is recovered from the No. 2 liquid outlet.
The recovered acid liquor is returned to step 1.1 for use in formulating the leaching agent.
Step 4, removing iron in the pretreatment liquid after regulating the pH value
Aerating the pretreatment liquid which enters the aeration tank after the pH value is regulated at 60-80 ℃ for 1-6 h to obtain slurry.
And filtering and washing the slurry to obtain an extraction stock solution and iron-removing slag.
Dehydrating the iron-removing slag for 2-5 hours at the temperature of 400-600 ℃ to obtain an iron phosphate product.
Step 5, purifying and enriching vanadium
Mixing the organic phase with the extracting stock solution according to the volume ratio of the organic phase to the extracting stock solution of 1:2-3, carrying out countercurrent forward extraction at the temperature of 10-30 ℃ and phase separation to obtain raffinate and a loaded organic phase.
Mixing the loaded organic phase with the stripping liquid according to the volume ratio of the loaded organic phase to the stripping liquid of (4-10) to 1, carrying out countercurrent stripping at the temperature of 10-30 ℃ and carrying out phase separation to obtain vanadium-rich liquid and organic-lean phase.
And (3) returning the raffinate to the step (3.2) and injecting the raffinate from the No. 2 liquid inlet.
The organic phase is a mixture of 10 to 25vol% of phosphoric acid extractant, 3 to 7vol% of phase regulator and 68 to 87vol% of sulfonated kerosene.
The back extraction liquid is sulfuric acid solution with the concentration of 6-10 vol%.
Step 6, regenerating the lean organic phase
Mixing the lean organic phase with the regenerant according to the volume ratio of (1-3) to 1, carrying out countercurrent regeneration at the temperature of 10-30 ℃ and separating phases to obtain regenerated liquid and regenerated organic phase.
And (5) returning the regenerated liquid to the step (5) to be used as a back extraction liquid.
And returning the regenerated organic phase to the step 5 for recycling as an organic phase.
The regenerant is 8-15 vol% sulfuric acid solution.
Step 7, precipitating vanadium and preparing vanadium pentoxide product
Mixing the vanadium-rich liquid and sodium chlorate uniformly according to the molar ratio of vanadium to chlorine of (3-6) to 1, and stirring and reacting for 1-3 hours at 50-70 ℃ to obtain the vanadium-rich liquid.
Mixing the vanadium-rich oxide solution and ammonium salt uniformly according to the molar ratio of vanadium to nitrogen of 1:2-6, regulating the pH value to 1.8-2.2 by sulfuric acid, stirring for 1-3 h at 75-95 ℃, filtering, and washing to obtain ammonium polyvanadate and vanadium precipitation mother liquor.
Deamination is carried out on the ammonium polyvanadate for 1-3 hours at the temperature of 450-550 ℃ to obtain a vanadium pentoxide product.
And (3) returning the vanadium precipitation mother liquor to the step (2) and mixing with the vanadium-containing pickle liquor.
The ammonium salt is one of ammonium chloride, ammonium sulfate and ammonia water.
The concentration of the concentrated sulfuric acid is more than or equal to 98wt%.
The fluorine-containing leaching aid is one or more of calcium fluoride, sodium fluoride and potassium fluoride.
The content of V 2O5 in the vanadium shale is 0.3-3wt%; the content of Fe 2O3 is 1-10wt%.
The phosphorus-containing reducing agent is more than one of sodium hypophosphite, sodium phosphite, hypophosphorous acid and phosphorous acid.
The stirring rotating speed is 100-500 r/min.
The aeration gas is air or oxygen.
The concentration of the sodium sulfate solution is 1-5 wt%.
The phosphoric acid extractant is di (2-ethylhexyl) phosphate or mono 2-ethylhexyl phosphate.
The phase regulator is sec-octanol or tributyl phosphate.
By adopting the method, compared with the prior art, the invention has the following positive effects:
1. Can effectively recycle F and residual acid in the acid leaching solution containing vanadium. In the leaching process, a fluorine-containing leaching aid is added, and the vanadium-containing acid leaching solution contains F with higher concentration. During the electrodialysis pH adjustment, about 30% of F will permeate the ion exchange membrane into the low acid solution; in the purification and enrichment process, as the phosphorus extractant hardly extracts F, the residual F in the vanadium-containing pickle liquor finally flows to raffinate; in the invention, raffinate is used for replacing water in the electrodialysis process in the subsequent circulation process, and the obtained recovered acid liquor can be returned to the leaching process, so that F can be recycled in the shale vanadium extraction process, F and residual acid in the vanadium-containing pickling solution are effectively utilized, and the addition amount of the subsequent fluorine-containing leaching aid and sulfuric acid is reduced.
2. The consumption of the medicament is small, and the medicament utilization rate is high. According to the invention, after Fe 3+ in the vanadium-containing pickle liquor is reduced to Fe 2+ by the phosphorus-containing reducing agent, the vanadium-containing pickle liquor is adjusted to a proper pH value by using an electrodialysis method instead of an alkali neutralization method, then Fe 2+ is oxidized to Fe 3+ by introducing air or oxygen, fe 3+ and PO 4 3- are combined to form a precipitate, and finally, iron in the vanadium-containing pickle liquor can be removed by solid-liquid separation, so that the pH value is not required to be adjusted by repeated acid-base neutralization, an iron extracting agent is not required, and a large amount of neutralizing agents and extracting agents are saved. The superfluous sulfuric acid and fluorine-containing leaching aid added in the leaching process are recycled in the form of recovered acid liquor finally; the PO 4 3- content in the vanadium-containing pickle liquor obtained by the method is far less than the Fe 3+ content, and the subsequently added phosphorus-containing reducing agent complements the PO 4 3- which is lacking in the vanadium-containing pickle liquor, and is recovered in the form of FePO 4 products, so that the medicament consumption is small and the medicament utilization rate is high.
3. Can recycle the iron element in the acid leaching solution containing vanadium. On the one hand, as the pH value is in the range of 1.3-2.0, almost only FePO 4 in the vanadium-containing pickle liquor can be separated out in the form of hydrated precipitate; on the other hand, the method does not use alkali for neutralization to adjust the pH value, no neutralization slag is generated, and FePO 4 precipitate cannot be mixed with the neutralization slag to be separated and recycled. Therefore, after the vanadium-containing pickle liquor is treated by the method, the obtained iron-removing slag has less impurities, single phase and easy solid-liquid separation, and the prepared iron phosphate product meets the requirements of iron phosphate industry standard (HG/T4701-2021) for batteries, so that the iron in the vanadium-containing pickle liquor has higher recovery value.
4. No neutralization slag is generated, and the environment is friendly. After the vanadium-containing acid leaching solution is treated by the method provided by the invention, the obtained iron-removing slag can be recycled for preparing battery-grade ferric phosphate, the obtained recycled acid liquor can be prepared into a leaching agent for recycling in the leaching process of vanadium shale, and no neutralizing agent is needed to be added in the method provided by the invention, and no neutralizing slag is generated, so that source emission reduction is realized, and the method is environment-friendly.
5. The recovery liquid and the washing liquid can be recycled, and vanadium loss is avoided. Firstly, after the acid leaching solution containing vanadium is treated by the method, the retention rate of vanadium is measured to be more than 95wt%, so that the subsequent purification and enrichment are not influenced; second, although about 5wt% of the vanadium in the pickling solution will pass through the ion exchange membrane into the recovered acid solution, which is recycled for the pickling process, so that there is no loss of this vanadium; thirdly, no neutralization slag is generated in the treatment process, so that vanadium loss caused by the entrainment adsorption of a large amount of neutralization slag is avoided.
Therefore, the invention has the characteristics of effectively recovering vanadium and iron in the vanadium shale, easily recovering residual sulfuric acid and fluorine-containing leaching aid in the vanadium-containing acid leaching solution, no generation of neutralization slag, high medicament utilization rate, small medicament consumption and no vanadium loss.
Drawings
Fig. 1 is a schematic structural diagram of a device for adjusting the pH of a pretreatment liquid according to the present invention.
Detailed Description
The invention is further described in connection with the drawings and the detailed description which follow, without limiting the scope of the invention.
A method for recovering vanadium and iron from vanadium shale. The method in the specific embodiment comprises the following steps:
Step 1, directly leaching vanadium shale
Step 1.1, preparing a leaching agent
And uniformly mixing concentrated sulfuric acid and water or concentrated sulfuric acid, water and recovered acid liquid according to the volume fraction of sulfuric acid of 10-40%, thereby obtaining the leaching agent.
Step 1.2, preparing leached powder
Crushing the vanadium-containing shale, and ball-milling until the particle size is smaller than 74 mu m and 40-70 wt% of the shale is obtained.
The sodium chlorate, the fluorine-containing leaching aid and the raw ore powder are uniformly mixed according to the mass ratio of (0.01-0.05) to (0.01-0.1) to 1.
Step 1.3, direct acid leaching
And mixing the leaching agent and the leaching powder according to the liquid-solid ratio of 0.8-2.5L/kg to obtain leaching slurry.
Stirring the leached slurry for 6-24 hours at the temperature of 80-98 ℃, filtering, and washing to obtain the vanadium-containing pickle liquor and leached slag.
Step2, reduction of the vanadium-containing pickle liquor
Adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor according to the mol ratio of phosphorus to iron in the pretreatment liquor of (1-5) to 1, and stirring for 1-6 h at the temperature of 60-100 ℃ to prepare the pretreatment liquor.
Step3, adjusting the pH of the pretreatment liquid
Step 3.1, device for adjusting pH of pretreatment liquid
As shown in fig. 1, the device for adjusting the pH of the pretreatment liquid includes: cathode, anode, membrane stack and DC power supply.
The cathode is connected with the negative electrode of the direct current power supply, the anode is connected with the positive electrode of the direct current power supply, and the cathode and the anode are correspondingly arranged on the right side and the left side of the membrane stack.
The membrane stack sequentially comprises a 1 st anion exchange membrane, a 1 st cation exchange membrane, a 2 nd anion exchange membrane, a 2 nd cation exchange membrane, a 3 rd anion exchange membrane, … … th anion exchange membrane, an nth cation exchange membrane and an n+1th anion exchange membrane from the anode to the cathode; and n is a positive integer of 10-1000.
As shown in fig. 1, from the anode to the cathode: the gap between the anode and the 1 st anion exchange membrane forms an anode electrode chamber, the gap between the 1 st anion exchange membrane and the 1 st cation exchange membrane forms a 1-stage regulating chamber, the gap between the 1 st cation exchange membrane and the 2 nd anion exchange membrane forms an n-stage recovery acid chamber, the gap between the 2 nd anion exchange membrane and the 2 nd cation exchange membrane forms a 2-stage regulating chamber, the gap between the 2 nd cation exchange membrane and the 3 rd anion exchange membrane forms an n-1 stage recovery acid chamber, … …, and so on, the gap between the n-1 st anion exchange membrane and the n-1 st cation exchange membrane forms an n-1 stage regulating chamber, the gap between the n-1 st cation exchange membrane and the n anion exchange membrane forms an n-stage recovery acid chamber, the gap between the n cation exchange membrane and the n+1 th anion exchange membrane forms a 1-stage recovery acid chamber, and the gap between the n+1 th anion exchange membrane and the cathode electrode forms a cathode electrode chamber.
As shown in fig. 1, the lower part of the 1-level adjusting chamber, the lower part of the 2-level adjusting chamber, the lower part of the 3-level adjusting chamber, … … and the lower part of the n-level adjusting chamber are respectively communicated with the No. 1 liquid inlet, and the upper part of the 1-level adjusting chamber, the upper part of the 2-level adjusting chamber, the upper part of the 3-level adjusting chamber, … … and the upper part of the n-level adjusting chamber are respectively communicated with the No. 1 liquid outlet.
As shown in fig. 1, the lower part of the 1-level recovery acid chamber, the lower part of the 2-level recovery acid chamber, the lower part of the 3-level recovery acid chamber, … … and the lower part of the n-level recovery acid chamber are respectively communicated with the No. 2 liquid inlet, and the upper part of the 1-level recovery acid chamber, the upper part of the 2-level recovery acid chamber, the upper part of the 3-level recovery acid chamber, … … and the upper part of the n-level recovery acid chamber are respectively communicated with the No. 2 liquid outlet.
As shown in fig. 1, the anode electrode chamber, the 1-stage adjusting chamber, the n-stage acid recycling chamber, the 2-stage adjusting chamber, the n-1-stage acid recycling chamber, … …, the n-1-stage adjusting chamber, the 2-stage acid recycling chamber, the n-stage adjusting chamber, the 1-stage acid recycling chamber, the cathode electrode chamber and the direct current power supply form a series loop in a working state to obtain a device for adjusting the pH of the pretreatment liquid.
Step 3.2, adjusting the pH value of the pretreatment liquid
And (3) respectively injecting sodium sulfate solution into an anode electrode chamber and a cathode electrode chamber of the device for adjusting the pH of the pretreatment liquid, injecting the pretreatment liquid prepared in the step (2) from a No.1 liquid inlet of the device for adjusting the pH of the pretreatment liquid, and injecting water or raffinate from a No. 2 liquid inlet of the device for adjusting the pH of the pretreatment liquid.
And switching on a direct current power supply of the device for regulating the pH value of the pretreatment liquid.
The pretreatment liquid injected from the No. 1 liquid inlet flows through a 1-level adjusting chamber, a 2-level adjusting chamber, a 3-level adjusting chamber, … …, an n-1-level adjusting chamber and an n-level adjusting chamber simultaneously, then flows out from a No. 1 liquid outlet and returns to the No. 1 liquid inlet, and the pretreatment liquid is circularly reciprocated until the pH value of the pretreatment liquid after the pH value is adjusted, which flows out from the No. 1 liquid outlet, is 1.3-2.0, and then enters the aeration tank in the step 3.
The water injected from the No. 2 liquid inlet flows through the 1-grade recovery acid chamber, the 2-grade recovery acid chamber, the 3-grade recovery acid chamber, … …, the n-1 grade recovery acid chamber and the n-grade recovery acid chamber at the same time, and then flows out from the No. 2 liquid outlet and returns to the No. 2 liquid inlet, so that the water is circularly reciprocated until the pretreatment liquid flows out from the No. 1 liquid outlet after the pH value is regulated and enters the aeration tank in the step 3, and then the acid liquid is recovered from the No. 2 liquid outlet.
The recovered acid liquor is returned to step 1.1 for use in formulating the leaching agent.
Step 4, removing iron in the pretreatment liquid after regulating the pH value
Aerating the pretreatment liquid which enters the aeration tank after the pH value is regulated at 60-80 ℃ for 1-6 h to obtain slurry.
And filtering and washing the slurry to obtain an extraction stock solution and iron-removing slag.
Dehydrating the iron-removing slag for 2-5 hours at the temperature of 400-600 ℃ to obtain an iron phosphate product.
Step 5, purifying and enriching vanadium
Mixing the organic phase with the extracting stock solution according to the volume ratio of the organic phase to the extracting stock solution of 1:2-3, carrying out countercurrent forward extraction at the temperature of 10-30 ℃ and phase separation to obtain raffinate and a loaded organic phase.
Mixing the loaded organic phase with the stripping liquid according to the volume ratio of the loaded organic phase to the stripping liquid of (4-10) to 1, carrying out countercurrent stripping at the temperature of 10-30 ℃ and carrying out phase separation to obtain vanadium-rich liquid and organic-lean phase.
And (3) returning the raffinate to the step (3.2) and injecting the raffinate from the No. 2 liquid inlet.
The organic phase is a mixture of 10 to 25vol% of phosphoric acid extractant, 3 to 7vol% of phase regulator and 68 to 87vol% of sulfonated kerosene.
The back extraction liquid is sulfuric acid solution with the concentration of 6-10 vol%.
Step 6, regenerating the lean organic phase
Mixing the lean organic phase with the regenerant according to the volume ratio of (1-3) to 1, carrying out countercurrent regeneration at the temperature of 10-30 ℃ and separating phases to obtain regenerated liquid and regenerated organic phase.
And (5) returning the regenerated liquid to the step (5) to be used as a back extraction liquid.
And returning the regenerated organic phase to the step 5 for recycling as an organic phase.
The regenerant is 8-15 vol% sulfuric acid solution.
Step 7, precipitating vanadium and preparing vanadium pentoxide product
Mixing the vanadium-rich liquid and sodium chlorate uniformly according to the molar ratio of vanadium to chlorine of (3-6) to 1, and stirring and reacting for 1-3 hours at 50-70 ℃ to obtain the vanadium-rich liquid.
Mixing the vanadium-rich oxide solution and ammonium salt uniformly according to the molar ratio of vanadium to nitrogen of 1:2-6, regulating the pH value to 1.8-2.2 by sulfuric acid, stirring for 1-3 h at 75-95 ℃, filtering, and washing to obtain ammonium polyvanadate and vanadium precipitation mother liquor.
Deamination is carried out on the ammonium polyvanadate for 1-3 hours at the temperature of 450-550 ℃ to obtain a vanadium pentoxide product.
And (3) returning the vanadium precipitation mother liquor to the step (2) and mixing with the vanadium-containing pickle liquor.
The ammonium salt is one of ammonium chloride, ammonium sulfate and ammonia water.
The concentration of the concentrated sulfuric acid is more than or equal to 98wt%.
The fluorine-containing leaching aid is one or more of calcium fluoride, sodium fluoride and potassium fluoride.
The content of V 2O5 in the vanadium shale is 0.3-3wt%; the content of Fe 2O3 is 1-10wt%.
The phosphorus-containing reducing agent is more than one of sodium hypophosphite, sodium phosphite, hypophosphorous acid and phosphorous acid.
The stirring rotating speed is 100-500 r/min.
The aeration gas is air or oxygen.
The concentration of the sodium sulfate solution is 1-5 wt%.
The phosphoric acid extractant is di (2-ethylhexyl) phosphate or mono 2-ethylhexyl phosphate.
The phase regulator is sec-octanol or tributyl phosphate.
Example 1
A method for recovering vanadium and iron from vanadium shale. The method in the specific embodiment comprises the following steps:
Step 1, directly leaching vanadium shale
Step 1.1, preparing a leaching agent
And uniformly mixing concentrated sulfuric acid and water according to the volume fraction of sulfuric acid of 10%, thus obtaining the leaching agent.
Step 1.2, preparing leached powder
Crushing the vanadium-containing shale, and ball-milling until the particle size is smaller than 74 mu m and 40wt% of the shale is obtained.
And uniformly mixing the sodium chlorate, the fluorine-containing leaching aid and the raw ore powder according to the mass ratio of the sodium chlorate to the fluorine-containing leaching aid to the raw ore powder of 0.03:0.07:1 to obtain leaching powder.
Step 1.3, direct acid leaching
And mixing the leaching agent and the leaching powder according to the liquid-solid ratio of 2.5L/kg to obtain leaching slurry.
Stirring the leaching slurry for 14h at 92 ℃, filtering, and washing to obtain vanadium-containing pickle liquor and leaching residues.
Step2, reduction of the vanadium-containing pickle liquor
Adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor according to the mol ratio of phosphorus to iron in the pretreatment liquor of 1:1, and stirring for 6 hours at the temperature of 100 ℃ to prepare the pretreatment liquor.
Step3, adjusting the pH of the pretreatment liquid
Step 3.1, device for adjusting pH of pretreatment liquid
As shown in fig. 1, the device for adjusting the pH of the pretreatment liquid includes: cathode, anode, membrane stack and DC power supply.
The cathode is connected with the negative electrode of the direct current power supply, the anode is connected with the positive electrode of the direct current power supply, and the cathode and the anode are correspondingly arranged on the right side and the left side of the membrane stack.
The membrane stack sequentially comprises a1 st anion exchange membrane, a1 st cation exchange membrane, a2 nd anion exchange membrane, a2 nd cation exchange membrane, a3 rd anion exchange membrane, … … th anion exchange membrane, an nth cation exchange membrane and an n+1th anion exchange membrane from the anode to the cathode; and n is a positive integer of 10.
As shown in fig. 1, from the anode to the cathode: the gap between the anode and the 1 st anion exchange membrane forms an anode electrode chamber, the gap between the 1 st anion exchange membrane and the 1 st cation exchange membrane forms a 1-stage regulating chamber, the gap between the 1 st cation exchange membrane and the 2 nd anion exchange membrane forms an n-stage recovery acid chamber, the gap between the 2 nd anion exchange membrane and the 2 nd cation exchange membrane forms a 2-stage regulating chamber, the gap between the 2 nd cation exchange membrane and the 3 rd anion exchange membrane forms an n-1 stage recovery acid chamber, … …, and so on, the gap between the n-1 st anion exchange membrane and the n-1 st cation exchange membrane forms an n-1 stage regulating chamber, the gap between the n-1 st cation exchange membrane and the n anion exchange membrane forms an n-stage recovery acid chamber, the gap between the n cation exchange membrane and the n+1 th anion exchange membrane forms a 1-stage recovery acid chamber, and the gap between the n+1 th anion exchange membrane and the cathode electrode forms a cathode electrode chamber.
As shown in fig. 1, the lower part of the 1-level adjusting chamber, the lower part of the 2-level adjusting chamber, the lower part of the 3-level adjusting chamber, … … and the lower part of the n-level adjusting chamber are respectively communicated with the No. 1 liquid inlet, and the upper part of the 1-level adjusting chamber, the upper part of the 2-level adjusting chamber, the upper part of the 3-level adjusting chamber, … … and the upper part of the n-level adjusting chamber are respectively communicated with the No. 1 liquid outlet.
As shown in fig. 1, the lower part of the 1-level recovery acid chamber, the lower part of the 2-level recovery acid chamber, the lower part of the 3-level recovery acid chamber, … … and the lower part of the n-level recovery acid chamber are respectively communicated with the No. 2 liquid inlet, and the upper part of the 1-level recovery acid chamber, the upper part of the 2-level recovery acid chamber, the upper part of the 3-level recovery acid chamber, … … and the upper part of the n-level recovery acid chamber are respectively communicated with the No. 2 liquid outlet.
As shown in fig. 1, the anode electrode chamber, the 1-stage adjusting chamber, the n-stage acid recycling chamber, the 2-stage adjusting chamber, the n-1-stage acid recycling chamber, … …, the n-1-stage adjusting chamber, the 2-stage acid recycling chamber, the n-stage adjusting chamber, the 1-stage acid recycling chamber, the cathode electrode chamber and the direct current power supply form a series loop in a working state to obtain a device for adjusting the pH of the pretreatment liquid.
Step 3.2, adjusting the pH value of the pretreatment liquid
And (3) respectively injecting sodium sulfate solution into an anode electrode chamber and a cathode electrode chamber of the device for adjusting the pH of the pretreatment liquid, injecting the pretreatment liquid prepared in the step (2) from a No.1 liquid inlet of the device for adjusting the pH of the pretreatment liquid, and injecting water or raffinate from a No. 2 liquid inlet of the device for adjusting the pH of the pretreatment liquid.
And switching on a direct current power supply of the device for regulating the pH value of the pretreatment liquid.
The pretreatment liquid injected from the No. 1 liquid inlet flows through a 1-level adjusting chamber, a 2-level adjusting chamber, a 3-level adjusting chamber, … …, an n-1-level adjusting chamber and an n-level adjusting chamber simultaneously, then flows out from a No. 1 liquid outlet and returns to the No. 1 liquid inlet, and the pretreatment liquid is circularly reciprocated until the pH value of the pretreatment liquid after the pH value is adjusted, which flows out from the No. 1 liquid outlet, is 1.8, and then enters the aeration tank in the step 3.
The water injected from the No. 2 liquid inlet flows through the 1-grade recovery acid chamber, the 2-grade recovery acid chamber, the 3-grade recovery acid chamber, … …, the n-1 grade recovery acid chamber and the n-grade recovery acid chamber at the same time, and then flows out from the No. 2 liquid outlet and returns to the No. 2 liquid inlet, so that the water is circularly reciprocated until the pretreatment liquid flows out from the No. 1 liquid outlet after the pH value is regulated and enters the aeration tank in the step 3, and then the acid liquid is recovered from the No. 2 liquid outlet.
The recovered acid liquor is returned to step 1.1 for use in formulating the leaching agent.
Step 4, removing iron in the pretreatment liquid after regulating the pH value
Aerating the pretreatment liquid which enters the aeration tank after the pH value is regulated for 4 hours at the temperature of 60 ℃ to obtain slurry.
And filtering and washing the slurry to obtain an extraction stock solution and iron-removing slag.
And dehydrating the iron-removing slag for 5 hours at 400 ℃ to obtain an iron phosphate product.
Step 5, purifying and enriching vanadium
Mixing the organic phase with the extraction stock solution according to the volume ratio of the organic phase to the extraction stock solution of 1:2, carrying out countercurrent forward extraction at the temperature of 10 ℃, and carrying out phase separation to obtain raffinate and a loaded organic phase.
Mixing the loaded organic phase with the stripping liquid according to the volume ratio of the loaded organic phase to the stripping liquid of 10:1, carrying out countercurrent stripping at the temperature of 10-30 ℃, and carrying out phase separation to obtain vanadium-rich liquid and organic-lean phase.
And (3) returning the raffinate to the step (3.2) and injecting the raffinate from the No. 2 liquid inlet.
The organic phase was a mixture of 10vol% phosphoric acid extractant, 3vol% phase modifier and 87vol% sulfonated kerosene.
The strip liquor is a sulfuric acid solution with the concentration of 6 vol%.
Step 6, regenerating the lean organic phase
Mixing the lean organic phase with the regenerant according to the volume ratio of 1:1, carrying out countercurrent regeneration at 10 ℃, and carrying out phase separation to obtain regenerated liquid and regenerated organic phase.
And (5) returning the regenerated liquid to the step (5) to be used as a back extraction liquid.
And returning the regenerated organic phase to the step 5 for recycling as an organic phase.
The regenerant was an 8vol% sulfuric acid solution.
Step 7, precipitating vanadium and preparing vanadium pentoxide product
Mixing the vanadium-rich liquid and sodium chlorate uniformly according to the molar ratio of vanadium to chlorine of 3:1, and stirring and reacting for 1h at 50 ℃ to obtain the vanadium-rich liquid.
Mixing the vanadium-rich oxide solution and ammonium salt uniformly according to the molar ratio of vanadium to nitrogen of 1:2, regulating the pH value to 1.8 by sulfuric acid, stirring for 1h at 75 ℃, filtering, and washing to obtain ammonium polyvanadate and vanadium precipitation mother liquor.
Deamination is carried out on the ammonium polyvanadate for 3 hours at the temperature of 450 ℃ to obtain a vanadium pentoxide product.
And (3) returning the vanadium precipitation mother liquor to the step (2) and mixing with the vanadium-containing pickle liquor.
The ammonium salt is ammonium chloride.
The concentration of the concentrated sulfuric acid is more than or equal to 98wt%.
The fluorine-containing leaching aid is calcium fluoride.
The content of V 2O5 in the vanadium shale is 0.3wt%; the content of Fe 2O3 was 1wt%.
The phosphorus-containing reducing agent is sodium phosphite.
The stirring rotating speed is 100r/min.
The aeration gas is air.
The concentration of the sodium sulfate solution was 4wt%.
The phosphoric acid extractant is di (2-ethylhexyl) phosphate.
The phase modifier is sec-octanol.
Example 2
A method for recovering vanadium and iron from vanadium shale. The method in the specific embodiment comprises the following steps:
Step 1, directly leaching vanadium shale
Step 1.1, preparing a leaching agent
Uniformly mixing concentrated sulfuric acid, water and recovered acid liquor according to the volume fraction of sulfuric acid of 20%, so as to obtain a leaching agent; the concentration of concentrated sulfuric acid was the same as in example 1.
Step 1.2, preparing leached powder
Crushing the vanadium-containing shale, and ball-milling until the particle size is less than 74 mu m and 50wt% of the shale is obtained.
And uniformly mixing the sodium chlorate, the fluorine-containing leaching aid and the raw ore powder according to the mass ratio of the sodium chlorate to the fluorine-containing leaching aid to the raw ore powder of 0.02:0.1:1 to obtain leaching powder.
Step 1.3, direct acid leaching
And mixing the leaching agent and the leaching powder according to the liquid-solid ratio of 2L/kg to obtain leaching slurry.
Stirring the leaching slurry for 8 hours at the temperature of 95 ℃, filtering, and washing to obtain vanadium-containing pickle liquor and leaching residues.
Step2, reduction of the vanadium-containing pickle liquor
Adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor according to the mol ratio of phosphorus to iron in the pretreatment liquor of 2:1, and stirring for 4 hours at the temperature of 85 ℃ to prepare the pretreatment liquor.
Step3, adjusting the pH of the pretreatment liquid
Step 3.1, device for adjusting pH of pretreatment liquid
The apparatus for adjusting the pH of the pretreatment liquid used in this step was the same as in example 1 except for the following parameters:
And n is a positive integer of 12.
Step 3.2, adjusting the pH value of the pretreatment liquid
And (3) entering an aeration tank in the step (3) until the pH value of the pretreatment liquid after the pH value is regulated flowing out of the liquid outlet No. 1 is 2.0.
Step 4, removing iron in the pretreatment liquid after regulating the pH value
Aerating the pretreatment liquid which enters the aeration tank after the pH value is regulated for 1h at 65 ℃ to obtain slurry.
And filtering and washing the slurry to obtain an extraction stock solution and iron-removing slag.
And dehydrating the iron-removing slag for 3 hours at the temperature of 500 ℃ to obtain an iron phosphate product.
Step 5, purifying and enriching vanadium
Mixing the organic phase with the extraction stock solution according to the volume ratio of the organic phase to the extraction stock solution of 1:2.1, carrying out countercurrent forward extraction at 25 ℃, and carrying out phase separation to obtain raffinate and a loaded organic phase.
Mixing the loaded organic phase with the stripping liquid according to the volume ratio of the loaded organic phase to the stripping liquid of 8:1, carrying out countercurrent stripping at 25 ℃, and carrying out phase separation to obtain vanadium-rich liquid and organic-poor phase.
And (3) returning the raffinate to the step (3.2) and injecting the raffinate from the No. 2 liquid inlet.
The organic phase was a mixture of 15vol% phosphoric acid extractant, 4vol% phase modifier and 81vol% sulfonated kerosene.
The strip liquor is 8vol% sulfuric acid solution.
Step 6, regenerating the lean organic phase
Mixing the lean organic phase with the regenerant according to the volume ratio of 2:1, carrying out countercurrent regeneration at 25 ℃, and carrying out phase separation to obtain regenerated liquid and regenerated organic phase.
And (5) returning the regenerated liquid to the step (5) to be used as a back extraction liquid.
And returning the regenerated organic phase to the step 5 for recycling as an organic phase.
The regenerant was an 11vol% sulfuric acid solution.
Step 7, precipitating vanadium and preparing vanadium pentoxide product
Mixing the vanadium-rich liquid and sodium chlorate uniformly according to the molar ratio of vanadium to chlorine of 4:1, and stirring and reacting for 2 hours at 55 ℃ to obtain the vanadium-rich liquid.
Mixing the vanadium-rich oxide solution and ammonium salt uniformly according to the molar ratio of vanadium to nitrogen of 1:3, regulating the pH value to 2.0 by sulfuric acid, stirring for 1.5h at 90 ℃, filtering, and washing to obtain ammonium polyvanadate and vanadium precipitation mother liquor.
Deamination is carried out on the ammonium polyvanadate for 2.5 hours at 470 ℃ to obtain a vanadium pentoxide product.
And (3) returning the vanadium precipitation mother liquor to the step (2) and mixing with the vanadium-containing pickle liquor.
The ammonium salt is ammonium sulfate.
The fluorine-containing leaching aid is sodium fluoride.
The content of V 2O5 in the vanadium shale is 1.5wt%; the content of Fe 2O3 was 2wt%.
The phosphorus-containing reducing agent is phosphorous acid.
The stirring rotating speed is 200r/min.
The aeration gas is oxygen.
The concentration of the sodium sulfate solution was 1wt%.
The phosphoric acid extractant is 2-ethylhexyl phosphoric acid mono-2-ethylhexyl ester.
The phase modifier is sec-octanol.
Example 3
A method for recovering vanadium and iron from vanadium shale. The method in the specific embodiment comprises the following steps:
Step 1, directly leaching vanadium shale
Step 1.1, preparing a leaching agent
Uniformly mixing concentrated sulfuric acid, water and recovered acid liquor according to the volume fraction of sulfuric acid of 30%, so as to obtain a leaching agent; the concentration of concentrated sulfuric acid was the same as in example 1.
Step 1.2, preparing leached powder
Crushing the vanadium-containing shale, and ball-milling until the particle size is smaller than 74 mu m and 60wt% of the shale is obtained, thereby obtaining raw ore powder.
And uniformly mixing the sodium chlorate, the fluorine-containing leaching aid and the raw ore powder according to the mass ratio of the sodium chlorate to the fluorine-containing leaching aid to the raw ore powder of 0.01:0.01:1 to obtain leaching powder.
Step 1.3, direct acid leaching
And mixing the leaching agent and the leaching powder according to the liquid-solid ratio of 1.5L/kg to obtain leaching slurry.
Stirring the leached slurry for 12 hours at 90 ℃, filtering, and washing to obtain a vanadium-containing pickle liquor and leached slag.
Step2, reduction of the vanadium-containing pickle liquor
Adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor according to the mol ratio of phosphorus to iron in the pretreatment liquor of 5:1, and stirring for 3 hours at 75 ℃ to prepare the pretreatment liquor.
Step3, adjusting the pH of the pretreatment liquid
Step 3.1, device for adjusting pH of pretreatment liquid
The apparatus for adjusting the pH of the pretreatment liquid used in this step was the same as in example 1 except for the following parameters:
and n is a positive integer of 500.
Step 3.2, adjusting the pH value of the pretreatment liquid
And (3) entering an aeration tank in the step (3) until the pH value of the pretreatment liquid after the pH value is regulated flowing out of the liquid outlet No. 1 is 1.3.
Step 4, removing iron in the pretreatment liquid after regulating the pH value
Aerating the pretreatment liquid which enters the aeration tank after the pH value is regulated for 6 hours at the temperature of 75 ℃ to obtain slurry.
And filtering and washing the slurry to obtain an extraction stock solution and iron-removing slag.
And dehydrating the iron-removing slag for 2 hours at 600 ℃ to obtain an iron phosphate product.
Step 5, purifying and enriching vanadium
Mixing the organic phase with the extraction stock solution according to the volume ratio of the organic phase to the extraction stock solution of 1:3, carrying out countercurrent forward extraction at 20 ℃, and carrying out phase separation to obtain raffinate and a loaded organic phase.
Mixing the loaded organic phase with the stripping liquid according to the volume ratio of the loaded organic phase to the stripping liquid of 6:1, carrying out countercurrent stripping at 20 ℃, and carrying out phase separation to obtain vanadium-rich liquid and organic-poor phase.
And (3) returning the raffinate to the step (3.2) and injecting the raffinate from the No. 2 liquid inlet.
The organic phase was a mixture of 20vol% phosphoric acid extractant, 6vol% phase modifier and 74vol% sulfonated kerosene.
The strip liquor is a sulfuric acid solution with the concentration of 10 vol%.
Step 6, regenerating the lean organic phase
Mixing the lean organic phase with the regenerant according to the volume ratio of 3:1, carrying out countercurrent regeneration at 20 ℃, and carrying out phase separation to obtain regenerated liquid and regenerated organic phase.
And (5) returning the regenerated liquid to the step (5) to be used as a back extraction liquid.
And returning the regenerated organic phase to the step 5 for recycling as an organic phase.
The regenerant was a 14vol% sulfuric acid solution.
Step 7, precipitating vanadium and preparing vanadium pentoxide product
Mixing the vanadium-rich liquid and sodium chlorate uniformly according to the molar ratio of vanadium to chlorine of 5:1, and stirring and reacting for 3 hours at 60 ℃ to obtain the vanadium-rich liquid.
Mixing the vanadium-rich oxide solution and ammonium salt uniformly according to the molar ratio of vanadium to nitrogen of 1:4, regulating the pH value to 2.1 by sulfuric acid, stirring for 2 hours at 95 ℃, filtering, and washing to obtain ammonium polyvanadate and vanadium precipitation mother liquor.
Deamination is carried out on the ammonium polyvanadate for 2 hours at the temperature of 500 ℃ to obtain a vanadium pentoxide product.
And (3) returning the vanadium precipitation mother liquor to the step (2) and mixing with the vanadium-containing pickle liquor.
The ammonium salt is ammonia water.
The fluorine-containing leaching aid is potassium fluoride.
The content of V 2O5 in the vanadium shale is 0.7wt%; the content of Fe 2O3 was 3wt%.
The phosphorus-containing reducing agent is sodium hypophosphite.
The stirring rotating speed is 300r/min.
The aeration gas is air.
The concentration of the sodium sulfate solution was 5wt%.
The phosphoric acid extractant is di (2-ethylhexyl) phosphate.
The phase regulator is tributyl phosphate.
Example 4
A method for recovering vanadium and iron from vanadium shale. The method in the specific embodiment comprises the following steps:
Step 1, directly leaching vanadium shale
Step 1.1, preparing a leaching agent
Uniformly mixing concentrated sulfuric acid, water and recovered acid liquor according to the volume fraction of the sulfuric acid of 40%, so as to obtain a leaching agent; the concentration of concentrated sulfuric acid was the same as in example 1.
Step 1.2, preparing leached powder
Crushing the vanadium-containing shale, and ball-milling until the particle size is less than 74 mu m and 70wt% of the shale is obtained to obtain raw ore powder.
And uniformly mixing the sodium chlorate, the fluorine-containing leaching aid and the raw ore powder according to the mass ratio of the sodium chlorate to the fluorine-containing leaching aid to the raw ore powder of 0.025:0.04:1 to obtain leaching powder.
Step 1.3, direct acid leaching
And mixing the leaching agent and the leaching powder according to the liquid-solid ratio of 0.8L/kg to obtain leaching slurry.
Stirring the leached slurry for 6 hours at 98 ℃, filtering, and washing to obtain vanadium-containing pickle liquor and leached slag.
Step2, reduction of the vanadium-containing pickle liquor
Adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor according to the mol ratio of phosphorus to iron in the pretreatment liquor of 4:1, and stirring for 1h at 60 ℃ to prepare the pretreatment liquor.
Step3, adjusting the pH of the pretreatment liquid
Step 3.1, device for adjusting pH of pretreatment liquid
The apparatus for adjusting the pH of the pretreatment liquid used in this step was the same as in example 1 except for the following parameters:
and n is a positive integer of 1000.
Step 3.2, adjusting the pH value of the pretreatment liquid
And (3) entering an aeration tank in the step (3) until the pH value of the pretreatment liquid after the pH value is regulated flowing out of the liquid outlet No. 1 is 1.5.
Step 4, removing iron in the pretreatment liquid after regulating the pH value
Aerating the pretreatment liquid which enters the aeration tank after the pH value is regulated for 3 hours at the temperature of 80 ℃ to obtain slurry.
And filtering and washing the slurry to obtain an extraction stock solution and iron-removing slag.
And dehydrating the iron-removing slag for 4.5 hours at 420 ℃ to obtain an iron phosphate product.
Step 5, purifying and enriching vanadium
Mixing the organic phase with the extraction stock solution according to the volume ratio of the organic phase to the extraction stock solution of 1:2.9, carrying out countercurrent forward extraction at 30 ℃, and carrying out phase separation to obtain raffinate and a loaded organic phase.
Mixing the loaded organic phase with the stripping liquid according to the volume ratio of the loaded organic phase to the stripping liquid of 4:1, carrying out countercurrent stripping at 30 ℃, and carrying out phase separation to obtain vanadium-rich liquid and organic-poor phase.
And (3) returning the raffinate to the step (3.2) and injecting the raffinate from the No. 2 liquid inlet.
The organic phase was a mixture of 25vol% phosphoric acid extractant, 7vol% phase modifier and 68vol% sulfonated kerosene.
The strip liquor is a sulfuric acid solution with the concentration of 9 vol%.
Step 6, regenerating the lean organic phase
Mixing the lean organic phase with the regenerant according to the volume ratio of 2.5:1, carrying out countercurrent regeneration at 30 ℃, and separating phases to obtain regenerated liquid and regenerated organic phase.
And (5) returning the regenerated liquid to the step (5) to be used as a back extraction liquid.
And returning the regenerated organic phase to the step 5 for recycling as an organic phase.
The regenerant was a 12vol% sulfuric acid solution.
Step 7, precipitating vanadium and preparing vanadium pentoxide product
Mixing the vanadium-rich liquid and sodium chlorate uniformly according to the molar ratio of vanadium to chlorine of 6:1, and stirring and reacting for 1.5h at 65 ℃ to obtain the vanadium-rich liquid.
Mixing the vanadium-rich oxide solution and ammonium salt uniformly according to the molar ratio of vanadium to nitrogen of 1:5, regulating the pH value to 1.9 by sulfuric acid, stirring for 2.5h at 87 ℃, filtering, and washing to obtain ammonium polyvanadate and vanadium precipitation mother liquor.
Deamination is carried out on the ammonium polyvanadate for 1.5 hours at the temperature of 520 ℃ to obtain a vanadium pentoxide product.
And (3) returning the vanadium precipitation mother liquor to the step (2) and mixing with the vanadium-containing pickle liquor.
The ammonium salt is ammonium chloride.
The fluorine-containing leaching aid is a mixture of calcium fluoride and sodium fluoride.
The content of V 2O5 in the vanadium shale is 1.8wt%; the content of Fe 2O3 was 4wt%.
The phosphorus-containing reducing agent is hypophosphorous acid.
The stirring rotating speed is 400r/min.
The aeration gas is oxygen.
The concentration of the sodium sulfate solution was 3wt%.
The phosphoric acid extractant is 2-ethylhexyl phosphoric acid mono-2-ethylhexyl ester.
The phase regulator is tributyl phosphate.
Example 5
A method for recovering vanadium and iron from vanadium shale. The method in the specific embodiment comprises the following steps:
Step 1, directly leaching vanadium shale
Step 1.1, preparing a leaching agent
Uniformly mixing concentrated sulfuric acid, water and recovered acid liquor according to the volume fraction of the sulfuric acid of 25 percent to obtain a leaching agent; the concentration of concentrated sulfuric acid was the same as in example 1.
Step 1.2, preparing leached powder
Crushing the vanadium-containing shale, and ball-milling until the particle size is smaller than 74 mu m and accounts for 55wt%, thereby obtaining raw ore powder.
And uniformly mixing the sodium chlorate, the fluorine-containing leaching aid and the raw ore powder according to the mass ratio of the sodium chlorate to the fluorine-containing leaching aid to the raw ore powder of 0.05:0.05:1 to obtain leaching powder.
Step 1.3, direct acid leaching
And mixing the leaching agent and the leaching powder according to the liquid-solid ratio of 1.3L/kg to obtain leaching slurry.
Stirring the leaching slurry for 24 hours at the temperature of 80 ℃, filtering, and washing to obtain vanadium-containing pickle liquor and leaching residues.
Step2, reduction of the vanadium-containing pickle liquor
Adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor according to the mol ratio of phosphorus to iron in the pretreatment liquor of 3:1, and stirring for 5 hours at 90 ℃ to prepare the pretreatment liquor.
Step3, adjusting the pH of the pretreatment liquid
Step 3.1, device for adjusting pH of pretreatment liquid
The apparatus for adjusting the pH of the pretreatment liquid used in this step was the same as in example 1 except for the following parameters:
and n is a positive integer of 50.
Step 3.2, adjusting the pH value of the pretreatment liquid
And (3) entering an aeration tank in the step (3) until the pH value of the pretreatment liquid after the pH value is regulated flowing out of the liquid outlet No. 1 is 1.6.
Step 4, removing iron in the pretreatment liquid after regulating the pH value
Aerating the pretreatment liquid which enters the aeration tank after the pH value is regulated for 5 hours at 70 ℃ to obtain slurry.
And filtering and washing the slurry to obtain an extraction stock solution and iron-removing slag.
And dehydrating the iron-removing slag for 4 hours at the temperature of 450 ℃ to obtain an iron phosphate product.
Step 5, purifying and enriching vanadium
Mixing the organic phase with the extraction stock solution according to the volume ratio of the organic phase to the extraction stock solution of 1:2.5, carrying out countercurrent forward extraction at 15 ℃, and carrying out phase separation to obtain raffinate and a loaded organic phase.
Mixing the loaded organic phase with the stripping liquid according to the volume ratio of the loaded organic phase to the stripping liquid of 7:1, carrying out countercurrent stripping at 15 ℃, and carrying out phase separation to obtain vanadium-rich liquid and organic-poor phase.
And (3) returning the raffinate to the step (3.2) and injecting the raffinate from the No. 2 liquid inlet.
The organic phase was a mixture of 12vol% phosphoric acid extractant, 5vol% phase modifier and 83vol% sulfonated kerosene.
The strip liquor is a sulfuric acid solution with the concentration of 7 vol%.
Step 6, regenerating the lean organic phase
Mixing the lean organic phase with the regenerant according to the volume ratio of 1.5:1, carrying out countercurrent regeneration at 15 ℃, and separating phases to obtain regenerated liquid and regenerated organic phase.
And (5) returning the regenerated liquid to the step (5) to be used as a back extraction liquid.
And returning the regenerated organic phase to the step 5 for recycling as an organic phase.
The regenerant was a 9vol% sulfuric acid solution.
Step 7, precipitating vanadium and preparing vanadium pentoxide product
Mixing the vanadium-rich liquid and sodium chlorate uniformly according to the molar ratio of vanadium to chlorine of 3.5:1, and stirring and reacting for 2.5 hours at 70 ℃ to obtain the vanadium-rich liquid.
Mixing the vanadium-rich oxide solution and ammonium salt uniformly according to the molar ratio of vanadium to nitrogen of 1:6, regulating the pH value to 2.2 by sulfuric acid, stirring for 3 hours at 93 ℃, filtering, and washing to obtain ammonium polyvanadate and vanadium precipitation mother liquor.
Deamination is carried out on the ammonium polyvanadate for 1h at 550 ℃ to obtain a vanadium pentoxide product.
And (3) returning the vanadium precipitation mother liquor to the step (2) and mixing with the vanadium-containing pickle liquor.
The ammonium salt is ammonium sulfate.
The fluorine-containing leaching aid is a mixture of sodium fluoride and potassium fluoride.
The content of V 2O5 in the vanadium shale is 0.9wt%; the content of Fe 2O3 was 6wt%.
The phosphorus-containing reducing agent is a mixture of sodium phosphite and phosphorous acid.
The stirring rotation speed is 500r/min.
The aeration gas is air.
The concentration of the sodium sulfate solution was 1.5wt%.
The phosphoric acid extractant is di (2-ethylhexyl) phosphate.
The phase modifier is sec-octanol.
Example 6
A method for recovering vanadium and iron from vanadium shale. The method in the specific embodiment comprises the following steps:
Step 1, directly leaching vanadium shale
Step 1.1, preparing a leaching agent
Uniformly mixing concentrated sulfuric acid, water and recovered acid liquor according to the volume fraction of the sulfuric acid of 35%, so as to obtain a leaching agent; the concentration of concentrated sulfuric acid was the same as in example 1.
Step 1.2, preparing leached powder
Crushing the vanadium-containing shale, and ball-milling until the particle size is less than 74 mu m and 65wt% of the shale is obtained.
And uniformly mixing the sodium chlorate, the fluorine-containing leaching aid and the raw ore powder according to the mass ratio of the sodium chlorate to the fluorine-containing leaching aid to the raw ore powder of 0.04:0.09:1 to obtain leaching powder.
Step 1.3, direct acid leaching
And mixing the leaching agent and the leaching powder according to the liquid-solid ratio of 1L/kg to obtain leaching slurry.
Stirring the leaching slurry for 20 hours at the temperature of 85 ℃, filtering, and washing to obtain vanadium-containing pickle liquor and leaching residues.
Step2, reduction of the vanadium-containing pickle liquor
Adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor according to the mol ratio of phosphorus to iron in the pretreatment liquor of 2.5:1, and stirring for 2 hours at 70 ℃ to prepare the pretreatment liquor.
Step3, adjusting the pH of the pretreatment liquid
Step 3.1, device for adjusting pH of pretreatment liquid
The apparatus for adjusting the pH of the pretreatment liquid used in this step was the same as in example 1 except for the following parameters:
And n is a positive integer of 700.
Step 3.2, adjusting the pH value of the pretreatment liquid
And (3) entering an aeration tank in the step (3) until the pH value of the pretreatment liquid after the pH value is regulated flowing out of the liquid outlet No. 1 is 1.7.
Step 4, removing iron in the pretreatment liquid after regulating the pH value
Aerating the pretreatment liquid which enters the aeration tank after the pH value is regulated for 2 hours at the temperature of 72 ℃ to obtain slurry.
And filtering and washing the slurry to obtain an extraction stock solution and iron-removing slag.
Dehydrating the iron-removing slag for 3.5 hours at 480 ℃ to obtain an iron phosphate product.
Step 5, purifying and enriching vanadium
Mixing the organic phase with the extraction stock solution according to the volume ratio of the organic phase to the extraction stock solution of 1:2.6, carrying out countercurrent forward extraction at 22 ℃ and phase separation to obtain raffinate and a loaded organic phase.
Mixing the loaded organic phase with the stripping liquid according to the volume ratio of the loaded organic phase to the stripping liquid of 5:1, carrying out countercurrent stripping at 22 ℃, and carrying out phase separation to obtain vanadium-rich liquid and organic-poor phase.
And (3) returning the raffinate to the step (3.2) and injecting the raffinate from the No. 2 liquid inlet.
The organic phase was a mixture of 17vol% phosphoric acid extractant, 4vol% phase modifier and 79vol% sulfonated kerosene.
The strip liquor is 8vol% sulfuric acid solution.
Step 6, regenerating the lean organic phase
Mixing the lean organic phase with the regenerant according to the volume ratio of 2.8:1, carrying out countercurrent regeneration at 22 ℃ and phase separation to obtain regenerated liquid and regenerated organic phase.
And (5) returning the regenerated liquid to the step (5) to be used as a back extraction liquid.
And returning the regenerated organic phase to the step 5 for recycling as an organic phase.
The regenerant was a 10vol% sulfuric acid solution.
Step 7, precipitating vanadium and preparing vanadium pentoxide product
Mixing the vanadium-rich liquid and sodium chlorate uniformly according to the molar ratio of vanadium to chlorine of 4.5:1, and stirring and reacting for 1.2 hours at 58 ℃ to obtain the vanadium-rich liquid.
Mixing the oxidized vanadium-rich liquid and ammonium salt uniformly according to the molar ratio of vanadium to nitrogen of 1:2.5, regulating the pH value to 2.0 by sulfuric acid, stirring for 2.2 hours at 82 ℃, filtering, and washing to obtain ammonium polyvanadate and vanadium precipitation mother liquor.
Deamination is carried out on the ammonium polyvanadate for 2.2 hours at the temperature of 510 ℃ to obtain a vanadium pentoxide product.
And (3) returning the vanadium precipitation mother liquor to the step (2) and mixing with the vanadium-containing pickle liquor.
The ammonium salt is ammonia water.
The fluorine-containing leaching aid is a mixture of calcium fluoride, sodium fluoride and potassium fluoride.
The content of V 2O5 in the vanadium shale is 1wt%; the content of Fe 2O3 was 7wt%.
The phosphorus-containing reducing agent is a mixture of sodium hypophosphite and hypophosphorous acid.
The stirring rotation speed is 350r/min.
The aeration gas is oxygen.
The concentration of the sodium sulfate solution was 2wt%.
The phosphoric acid extractant is 2-ethylhexyl phosphoric acid mono-2-ethylhexyl ester.
The phase regulator is tributyl phosphate.
Example 7
A method for recovering vanadium and iron from vanadium shale. The method in the specific embodiment comprises the following steps:
Step 1, directly leaching vanadium shale
Step 1.1, preparing a leaching agent
Uniformly mixing concentrated sulfuric acid, water and recovered acid liquor according to the volume fraction of sulfuric acid of 15%, so as to obtain a leaching agent; the concentration of concentrated sulfuric acid was the same as in example 1.
Step 1.2, preparing leached powder
Crushing the vanadium-containing shale, and ball-milling until the particle size is less than 74 mu m and 45wt% of the shale is obtained.
And uniformly mixing the sodium chlorate, the fluorine-containing leaching aid and the raw ore powder according to the mass ratio of the sodium chlorate to the fluorine-containing leaching aid to the raw ore powder of 0.015:0.02:1 to obtain leaching powder.
Step 1.3, direct acid leaching
And mixing the leaching agent and the leaching powder according to the liquid-solid ratio of 1.2L/kg to obtain leaching slurry.
Stirring the leached slurry for 16 hours at 82 ℃, filtering, and washing to obtain vanadium-containing pickle liquor and leached slag.
Step2, reduction of the vanadium-containing pickle liquor
Adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor according to the mol ratio of phosphorus to iron in the pretreatment liquor of 1.2:1, and stirring for 1.5 hours at 65 ℃ to prepare the pretreatment liquor.
Step3, adjusting the pH of the pretreatment liquid
Step 3.1, device for adjusting pH of pretreatment liquid
The apparatus for adjusting the pH of the pretreatment liquid used in this step was the same as in example 1 except for the following parameters:
And n is a positive integer of 300.
Step 3.2, adjusting the pH value of the pretreatment liquid
And (3) entering an aeration tank in the step (3) until the pH value of the pretreatment liquid after the pH value is regulated flowing out of the liquid outlet No.1 is 1.4.
Step 4, removing iron in the pretreatment liquid after regulating the pH value
Aerating the pretreatment liquid which enters the aeration tank after the pH value is regulated for 1.5 hours at 63 ℃ to obtain slurry.
And filtering and washing the slurry to obtain an extraction stock solution and iron-removing slag.
And dehydrating the iron-removing slag for 2.5 hours at the temperature of 520 ℃ to obtain an iron phosphate product.
Step 5, purifying and enriching vanadium
Mixing the organic phase with the extraction stock solution according to the volume ratio of the organic phase to the extraction stock solution of 1:2.2, carrying out countercurrent forward extraction at the temperature of 16 ℃ and phase separation to obtain raffinate and a loaded organic phase.
Mixing the loaded organic phase with the stripping liquid according to the volume ratio of the loaded organic phase to the stripping liquid of 9:1, carrying out countercurrent stripping at the temperature of 16 ℃, and carrying out phase separation to obtain vanadium-rich liquid and organic-poor phase.
And (3) returning the raffinate to the step (3.2) and injecting the raffinate from the No. 2 liquid inlet.
The organic phase was a mixture of 22vol% phosphoric acid extractant, 3vol% phase modifier, and 75vol% sulfonated kerosene.
The strip liquor is a sulfuric acid solution with the concentration of 7 vol%.
Step 6, regenerating the lean organic phase
Mixing the lean organic phase with the regenerant according to the volume ratio of 2.3:1, carrying out countercurrent regeneration at 16 ℃, and separating phases to obtain regenerated liquid and regenerated organic phase.
And (5) returning the regenerated liquid to the step (5) to be used as a back extraction liquid.
And returning the regenerated organic phase to the step 5 for recycling as an organic phase.
The regenerant was an 11vol% sulfuric acid solution.
Step 7, precipitating vanadium and preparing vanadium pentoxide product
Mixing the vanadium-rich liquid and sodium chlorate uniformly according to the molar ratio of vanadium to chlorine of 5.5:1, and stirring and reacting for 1.4 hours at 63 ℃ to obtain the vanadium-rich liquid.
Mixing the oxidized vanadium-rich liquid and ammonium salt uniformly according to the molar ratio of vanadium to nitrogen of 1:3.5, regulating the pH value to 1.9 by sulfuric acid, stirring for 1.2h at 86 ℃, filtering, and washing to obtain ammonium polyvanadate and vanadium precipitation mother liquor.
Deamination is carried out on the ammonium polyvanadate for 1.8 hours at the temperature of 460 ℃ to obtain a vanadium pentoxide product.
And (3) returning the vanadium precipitation mother liquor to the step (2) and mixing with the vanadium-containing pickle liquor.
The ammonium salt is ammonium chloride.
The fluorine-containing leaching aid is a mixture of calcium fluoride and potassium fluoride.
The content of V 2O5 in the vanadium shale is 2.7wt%; the content of Fe 2O3 was 8wt%.
The phosphorus-containing reducing agent is a mixture of sodium hypophosphite, sodium phosphite and hypophosphorous acid.
The stirring rotating speed is 450r/min.
The aeration gas is air.
The concentration of the sodium sulfate solution was 2.5wt%.
The phosphoric acid extractant is di (2-ethylhexyl) phosphate.
The phase modifier is sec-octanol.
Example 8
A method for recovering vanadium and iron from vanadium shale. The method in the specific embodiment comprises the following steps:
Step 1, directly leaching vanadium shale
Step 1.1, preparing a leaching agent
Uniformly mixing concentrated sulfuric acid, water and recovered acid liquor according to the volume fraction of the sulfuric acid of 29 percent to obtain a leaching agent; the concentration of concentrated sulfuric acid was the same as in example 1.
Step 1.2, preparing leached powder
Crushing the vanadium-containing shale, and ball-milling until the particle size is less than 74 mu m and 42wt% of the shale is obtained to obtain raw ore powder.
And uniformly mixing the sodium chlorate, the fluorine-containing leaching aid and the raw ore powder according to the mass ratio of the sodium chlorate to the fluorine-containing leaching aid to the raw ore powder of 0.035:0.03:1 to obtain leaching powder.
Step 1.3, direct acid leaching
And mixing the leaching agent and the leaching powder according to the liquid-solid ratio of 1.7L/kg to obtain leaching slurry.
Stirring the leached slurry for 10 hours at 88 ℃, filtering, and washing to obtain vanadium-containing pickle liquor and leached slag.
Step2, reduction of the vanadium-containing pickle liquor
Adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor according to the mol ratio of phosphorus to iron in the pretreatment liquor of 3.5:1, and stirring for 3.5 hours at 80 ℃ to prepare the pretreatment liquor.
Step3, adjusting the pH of the pretreatment liquid
Step 3.1, device for adjusting pH of pretreatment liquid
The apparatus for adjusting the pH of the pretreatment liquid used in this step was the same as in example 1 except for the following parameters:
And n is a positive integer of 400.
Step 3.2, adjusting the pH value of the pretreatment liquid
And (3) entering an aeration tank in the step (3) until the pH value of the pretreatment liquid after the pH value is regulated flowing out of the liquid outlet No.1 is 1.9.
Step 4, removing iron in the pretreatment liquid after regulating the pH value
Aerating the pretreatment liquid which enters the aeration tank after the pH value is regulated for 2.5 hours at 68 ℃ to obtain slurry.
And filtering and washing the slurry to obtain an extraction stock solution and iron-removing slag.
And dehydrating the iron-removing slag for 3.6 hours at 550 ℃ to obtain an iron phosphate product.
Step 5, purifying and enriching vanadium
Mixing the organic phase with the extraction stock solution according to the volume ratio of the organic phase to the extraction stock solution of 1:2.3, carrying out countercurrent forward extraction at 18 ℃ and phase separation to obtain raffinate and a loaded organic phase.
Mixing the loaded organic phase with the stripping liquid according to the volume ratio of the loaded organic phase to the stripping liquid of 5.5:1, carrying out countercurrent stripping at 18 ℃, and carrying out phase separation to obtain vanadium-rich liquid and organic-lean phase.
And (3) returning the raffinate to the step (3.2) and injecting the raffinate from the No. 2 liquid inlet.
The organic phase was a mixture of 23vol% phosphoric acid extractant, 5vol% phase modifier and 72vol% sulfonated kerosene.
The strip liquor is a sulfuric acid solution with the concentration of 6 vol%.
Step 6, regenerating the lean organic phase
Mixing the lean organic phase with the regenerant according to the volume ratio of 1.8:1, carrying out countercurrent regeneration at 18 ℃, and carrying out phase separation to obtain regenerated liquid and regenerated organic phase.
And (5) returning the regenerated liquid to the step (5) to be used as a back extraction liquid.
And returning the regenerated organic phase to the step 5 for recycling as an organic phase.
The regenerant was a 9vol% sulfuric acid solution.
Step 7, precipitating vanadium and preparing vanadium pentoxide product
Mixing the vanadium-rich liquid and sodium chlorate uniformly according to the molar ratio of vanadium to chlorine of 3.2:1, and stirring and reacting for 1.8 hours at 67 ℃ to obtain the vanadium-rich liquid.
Mixing the oxidized vanadium-rich liquid and ammonium salt uniformly according to the molar ratio of vanadium to nitrogen of 1:4.5, regulating the pH value to 2.1 by sulfuric acid, stirring for 1.6h at 78 ℃, filtering, and washing to obtain ammonium polyvanadate and vanadium precipitation mother liquor.
Deamination is carried out on the ammonium polyvanadate for 2.3 hours at the temperature of 480 ℃ to obtain a vanadium pentoxide product.
And (3) returning the vanadium precipitation mother liquor to the step (2) and mixing with the vanadium-containing pickle liquor.
The ammonium salt is ammonium sulfate.
The fluorine-containing leaching aid is calcium fluoride.
The content of V 2O5 in the vanadium shale is 3wt%; the content of Fe 2O3 was 5wt%.
The phosphorus-containing reducing agent is a mixture of sodium hypophosphite, sodium phosphite and phosphorous acid.
The stirring rotation speed is 250r/min.
The aeration gas is oxygen.
The concentration of the sodium sulfate solution was 3.5wt%.
The phosphoric acid extractant is 2-ethylhexyl phosphoric acid mono-2-ethylhexyl ester.
The phase regulator is tributyl phosphate.
Example 9
A method for recovering vanadium and iron from vanadium shale. The method in the specific embodiment comprises the following steps:
Step 1, directly leaching vanadium shale
Step 1.1, preparing a leaching agent
Uniformly mixing concentrated sulfuric acid, water and recovered acid liquor according to the volume fraction of the sulfuric acid of 37 percent to obtain a leaching agent; the concentration of concentrated sulfuric acid was the same as in example 1.
Step 1.2, preparing leached powder
Crushing the vanadium-containing shale, and ball-milling until the particle size is less than 74 mu m and 58wt% of the shale is obtained to obtain raw ore powder.
And uniformly mixing the sodium chlorate, the fluorine-containing leaching aid and the raw ore powder according to the mass ratio of the sodium chlorate to the fluorine-containing leaching aid to the raw ore powder of 0.045:0.06:1 to obtain leaching powder.
Step 1.3, direct acid leaching
And mixing the leaching agent and the leaching powder according to the liquid-solid ratio of 2.1L/kg to obtain leaching slurry.
Stirring the leached slurry for 18 hours at 89 ℃, filtering, and washing to obtain vanadium-containing pickle liquor and leached slag.
Step2, reduction of the vanadium-containing pickle liquor
Adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor according to the mole ratio of phosphorus to iron in the pretreatment liquor of 4.3:1, and stirring for 4.5 hours at the temperature of 95 ℃ to prepare the pretreatment liquor.
Step3, adjusting the pH of the pretreatment liquid
Step 3.1, device for adjusting pH of pretreatment liquid
The apparatus for adjusting the pH of the pretreatment liquid used in this step was the same as in example 1 except for the following parameters:
and n is a positive integer of 900.
Step 3.2, adjusting the pH value of the pretreatment liquid
And (3) entering an aeration tank in the step (3) until the pH value of the pretreatment liquid after the pH value is regulated flowing out of the liquid outlet No.1 is 1.8.
Step 4, removing iron in the pretreatment liquid after regulating the pH value
Aerating the pretreatment liquid which enters the aeration tank after the pH value is regulated for 3.5 hours at the temperature of 77 ℃ to obtain slurry.
And filtering and washing the slurry to obtain an extraction stock solution and iron-removing slag.
And dehydrating the iron-removing slag for 2.2 hours at 570 ℃ to obtain an iron phosphate product.
Step 5, purifying and enriching vanadium
Mixing the organic phase with the extraction stock solution according to the volume ratio of the organic phase to the extraction stock solution of 1:2.8, carrying out countercurrent forward extraction at 26 ℃, and carrying out phase separation to obtain raffinate and a loaded organic phase.
Mixing the loaded organic phase with the stripping liquid according to the volume ratio of the loaded organic phase to the stripping liquid of 7.5:1, carrying out countercurrent stripping at 26 ℃, and carrying out phase separation to obtain vanadium-rich liquid and organic-lean phase.
And (3) returning the raffinate to the step (3.2) and injecting the raffinate from the No. 2 liquid inlet.
The organic phase was a mixture of 19vol% phosphoric acid extractant, 6vol% phase modifier, and 75vol% sulfonated kerosene.
The strip liquor is a sulfuric acid solution with the concentration of 9 vol%.
Step 6, regenerating the lean organic phase
Mixing the lean organic phase with the regenerant according to the volume ratio of 1.2:1, carrying out countercurrent regeneration at 26 ℃, and separating phases to obtain regenerated liquid and regenerated organic phase.
And (5) returning the regenerated liquid to the step (5) to be used as a back extraction liquid.
And returning the regenerated organic phase to the step 5 for recycling as an organic phase.
The regenerant was a 15vol% sulfuric acid solution.
Step 7, precipitating vanadium and preparing vanadium pentoxide product
Mixing the vanadium-rich liquid and sodium chlorate uniformly according to the molar ratio of vanadium to chlorine of 4.8:1, and stirring and reacting for 2.2 hours at 69 ℃ to obtain the vanadium-rich liquid.
Mixing the oxidized vanadium-rich liquid and ammonium salt uniformly according to the molar ratio of vanadium to nitrogen of 1:5.5, regulating the pH value to 1.8 by sulfuric acid, stirring for 2.6h at 81 ℃, filtering, and washing to obtain ammonium polyvanadate and vanadium precipitation mother liquor.
Deamination is carried out on the ammonium polyvanadate for 2.8 hours at the temperature of 490 ℃ to obtain a vanadium pentoxide product.
And (3) returning the vanadium precipitation mother liquor to the step (2) and mixing with the vanadium-containing pickle liquor.
The ammonium salt is ammonia water.
The fluorine-containing leaching aid is sodium fluoride.
The content of V 2O5 in the vanadium shale is 2.2wt%; the content of Fe 2O3 was 10wt%.
The phosphorus-containing reducing agent is a mixture of sodium hypophosphite, sodium phosphite, hypophosphorous acid and phosphorous acid.
The stirring rotating speed is 150r/min.
The aeration gas is air.
The concentration of the sodium sulfate solution was 4.5wt%.
The phosphoric acid extractant is di (2-ethylhexyl) phosphate.
The phase modifier is sec-octanol.
Compared with the prior art, the specific embodiment has the following positive effects:
1. Can effectively recycle F and residual acid in the acid leaching solution containing vanadium. Fluorine-containing leaching aid is added in the leaching process of the specific embodiment, and the vanadium-containing leaching solution contains high concentration of F. During the electrodialysis pH adjustment, about 30% of F will permeate the ion exchange membrane into the low acid solution; in the purification and enrichment process, as the phosphorus extractant hardly extracts F, the residual F in the vanadium-containing pickle liquor finally flows to raffinate; in the specific embodiment, raffinate is used for replacing water in the electrodialysis process in the subsequent circulation process, the obtained recovered acid liquor can be returned to the leaching process, so that F can be recycled in the shale vanadium extraction process, F and residual acid in the vanadium-containing pickling solution are effectively utilized, and the addition amount of the subsequent fluorine-containing leaching aid and sulfuric acid is reduced.
2. The consumption of the medicament is small, and the medicament utilization rate is high. In the specific embodiment, after Fe 3+ in the vanadium-containing pickle liquor is reduced to Fe 2+ by a phosphorus-containing reducing agent, the vanadium-containing pickle liquor is adjusted to a proper pH value by an electrodialysis method instead of an alkali neutralization method, then Fe 2+ is oxidized to Fe 3+ by introducing air or oxygen, fe 3+ and PO 4 3- are combined to form a precipitate, and finally, the iron in the vanadium-containing pickle liquor can be removed by solid-liquid separation, so that the pH value does not need to be adjusted by repeated acid-base neutralization and an iron extractant is not needed, and a large amount of neutralizing agents and extracting agents are saved. The superfluous sulfuric acid and fluorine-containing leaching aid added in the leaching process are recycled in the form of recovered acid liquor finally; the PO 4 3- content in the vanadium-containing pickle liquor obtained by the method in the specific embodiment is far less than the Fe 3+ content, and the subsequently added phosphorus-containing reducing agent complements the PO 4 3- which is lacking in the vanadium-containing pickle liquor, and is recovered in the form of FePO 4 products, so that the medicament consumption is small and the medicament utilization rate is high.
3. Can recycle the iron element in the acid leaching solution containing vanadium. On the one hand, as the pH value is in the range of 1.3-2.0, almost only FePO 4 in the vanadium-containing pickle liquor can be separated out in the form of hydrated precipitate; on the other hand, the method in the specific embodiment does not use alkali for neutralization to adjust the pH value, no neutralization slag is generated, and FePO 4 precipitate cannot be mixed with the neutralization slag to be separated and recycled. Therefore, after the vanadium-containing pickle liquor is treated by the method in the specific embodiment, the obtained iron-removing slag has less impurities, single phase and easy solid-liquid separation, and the prepared iron phosphate product meets the requirements of the iron phosphate industry standard (HG/T4701-2021) for batteries, so that the iron in the vanadium-containing pickle liquor has higher recovery value.
4. No neutralization slag is generated, and the environment is friendly. After the vanadium-containing pickle liquor is treated by the method in the specific embodiment, the obtained iron-removing slag can be recycled for preparing battery-grade ferric phosphate, the obtained recycled acid liquor can be prepared into a leaching agent for recycling in the leaching process of vanadium shale, and the method in the specific embodiment does not need to add any neutralizing agent and does not generate neutralizing slag, so that source emission reduction is realized, and the method is environment-friendly.
5. The recovery liquid and the washing liquid can be recycled, and vanadium loss is avoided. Firstly, after the vanadium-containing pickle liquor is treated by the method in the specific embodiment, the retention rate of vanadium is measured to be more than 95wt percent, and the subsequent purification and enrichment are not affected; second, although about 5wt% of the vanadium in the pickling solution will pass through the ion exchange membrane into the recovered acid solution, which is recycled for the pickling process, so that there is no loss of this vanadium; third, no neutralization slag is generated in the treatment process of the specific embodiment, and vanadium loss caused by the entrainment adsorption of a large amount of neutralization slag is avoided.
Therefore, the specific embodiment has the characteristics of capability of effectively recovering vanadium and iron in the vanadium shale, easiness in recovering residual sulfuric acid and fluorine-containing leaching aid in the vanadium-containing pickle liquor, no generation of neutralization residues, high medicament utilization rate, small medicament consumption and no vanadium loss.
Claims (10)
1. A method for recovering vanadium and iron from vanadium shale, which is characterized by comprising the following steps:
Step 1, directly leaching vanadium shale
Step 1.1, preparing a leaching agent
Mixing concentrated sulfuric acid and water or concentrated sulfuric acid, water and recovered acid liquid uniformly according to the volume fraction of sulfuric acid of 10-40% to obtain a leaching agent;
Step 1.2, preparing leached powder
Crushing the vanadium-containing shale, and ball-milling until the particle size is smaller than 74 mu m and accounts for 40-70 wt% to obtain raw ore powder;
uniformly mixing sodium chlorate, a fluorine-containing leaching aid and raw ore powder according to the mass ratio of (0.01-0.05) to (0.01-0.1) to 1 of the sodium chlorate to the fluorine-containing leaching aid to obtain leaching powder;
Step 1.3, direct acid leaching
Mixing the leaching agent and the leaching powder according to the liquid-solid ratio of 0.8-2.5L/kg to obtain leaching slurry;
Stirring the leached slurry for 6-24 hours at the temperature of 80-98 ℃, filtering, and washing to obtain a vanadium-containing pickle liquor and leached slag;
Step2, reduction of the vanadium-containing pickle liquor
Adding a phosphorus-containing reducing agent into the vanadium-containing pickle liquor according to the molar ratio of phosphorus to iron in the pretreatment liquor of (1-5) to 1, and stirring for 1-6 hours at the temperature of 60-100 ℃ to prepare the pretreatment liquor;
Step3, adjusting the pH of the pretreatment liquid
Step 3.1, device for adjusting pH of pretreatment liquid
The device for adjusting the pH value of the pretreatment liquid comprises: a cathode, an anode, a membrane stack and a direct current power supply;
the cathode is connected with the negative electrode of the direct current power supply, the anode is connected with the positive electrode of the direct current power supply, and the cathode and the anode are correspondingly arranged on the right side and the left side of the membrane stack;
The membrane stack sequentially comprises a 1 st anion exchange membrane, a 1 st cation exchange membrane, a 2 nd anion exchange membrane, a 2 nd cation exchange membrane, a 3 rd anion exchange membrane, … … th anion exchange membrane, an nth cation exchange membrane and an n+1th anion exchange membrane from the anode to the cathode; n is a positive integer of 10-1000;
From the anode to cathode direction: the gap between the anode and the 1 st anion exchange membrane forms an anode electrode chamber, the gap between the 1 st anion exchange membrane and the 1 st cation exchange membrane forms a 1-stage regulating chamber, the gap between the 1 st cation exchange membrane and the 2 nd anion exchange membrane forms an n-stage recovered acid chamber, the gap between the 2 nd anion exchange membrane and the 2 nd cation exchange membrane forms a 2-stage recovered acid chamber, the gap between the 2 nd cation exchange membrane and the 3 rd anion exchange membrane forms an n-1 stage recovered acid chamber, … … and so on, the gap between the n-1 st anion exchange membrane and the n-1 st cation exchange membrane forms an n-1 stage recovered acid chamber, the gap between the n-1 st cation exchange membrane and the n-th anion exchange membrane forms an n-stage recovered acid chamber, the gap between the n cation exchange membrane and the n+1 th anion exchange membrane forms a cathode electrode chamber;
the lower part of the 1-level adjusting chamber, the lower part of the 2-level adjusting chamber, the lower part of the 3-level adjusting chamber, … … and the lower part of the n-level adjusting chamber are respectively connected with a No.1 liquid inlet, and the upper part of the 1-level adjusting chamber, the upper part of the 2-level adjusting chamber, the upper part of the 3-level adjusting chamber, the upper part of the … … and the upper part of the n-level adjusting chamber are respectively connected with a No.1 liquid outlet;
The lower part of the 1-level recovery acid chamber, the lower part of the 2-level recovery acid chamber, the lower part of the 3-level recovery acid chamber, … … and the lower part of the n-level recovery acid chamber are respectively connected with a No. 2 liquid inlet, and the upper part of the 1-level recovery acid chamber, the upper part of the 2-level recovery acid chamber, the upper part of the 3-level recovery acid chamber, … … and the upper part of the n-level recovery acid chamber are respectively connected with a No. 2 liquid outlet;
Forming a serial loop of the anode electrode chamber, the 1-stage regulating chamber, the n-stage acid recovering chamber, the 2-stage regulating chamber, the n-1-stage acid recovering chamber, … …, the n-1-stage regulating chamber, the 2-stage acid recovering chamber, the n-stage regulating chamber, the 1-stage acid recovering chamber, the cathode electrode chamber and the direct current power supply in a working state to obtain a device for regulating the pH value of the pretreatment liquid;
Step 3.2, adjusting the pH value of the pretreatment liquid
Injecting sodium sulfate solution into an anode electrode chamber and a cathode electrode chamber of the device for adjusting the pH of the pretreatment liquid respectively, injecting the pretreatment liquid prepared in the step 2 from a No.1 liquid inlet of the device for adjusting the pH of the pretreatment liquid, and injecting water or raffinate from a No. 2 liquid inlet of the device for adjusting the pH of the pretreatment liquid;
Switching on a direct current power supply of a device for adjusting the pH of the pretreatment liquid;
The pretreatment liquid injected from the No. 1 liquid inlet flows through a 1-level adjusting chamber, a 2-level adjusting chamber, a 3-level adjusting chamber, … …, an n-1-level adjusting chamber and an n-level adjusting chamber at the same time, then flows out from a No. 1 liquid outlet and returns to the No. 1 liquid inlet, and the pretreatment liquid is circularly reciprocated until the pH value of the pretreatment liquid after the pH value is adjusted, which flows out from the No. 1 liquid outlet, is 1.3-2.0, and then enters the aeration tank in the step 3;
The water injected from the No. 2 liquid inlet flows through a 1-grade recovery acid chamber, a 2-grade recovery acid chamber, a 3-grade recovery acid chamber, … …, an n-1 grade recovery acid chamber and an n-grade recovery acid chamber at the same time, then flows out from a No. 2 liquid outlet and returns to the No. 2 liquid inlet, and the water is circularly reciprocated, until the pretreatment liquid flows into the aeration tank in the step 3 after the pH value is regulated, and then the acid liquid is recovered from the No. 2 liquid outlet;
the recovered acid liquor returns to the step 1.1 for preparing the leaching agent;
Step 4, removing iron in the pretreatment liquid after regulating the pH value
Aerating the pretreatment liquid which enters the aeration tank after the pH value is regulated at 60-80 ℃ for 1-6 h to obtain slurry;
Filtering and washing the slurry to obtain an extraction stock solution and iron-removing slag;
Dehydrating the iron-removing slag at 400-600 ℃ for 2-5 h to obtain an iron phosphate product;
Step 5, purifying and enriching vanadium
Mixing the organic phase with the extracting stock solution according to the volume ratio of the organic phase to the extracting stock solution of 1:2-3, carrying out countercurrent forward extraction at the temperature of 10-30 ℃ and phase separation to obtain raffinate and a loaded organic phase;
Mixing the loaded organic phase with the stripping liquid according to the volume ratio of the loaded organic phase to the stripping liquid of (4-10) to 1, carrying out countercurrent stripping at the temperature of 10-30 ℃, and carrying out phase separation to obtain vanadium-rich liquid and organic-lean phase;
returning the raffinate to the step 3.2, and injecting the raffinate from the No. 2 liquid inlet;
the organic phase is a mixture of 10-25 vol% of phosphoric acid extractant, 3-7 vol% of phase regulator and 68-87 vol% of sulfonated kerosene;
The back extraction liquid is sulfuric acid solution with the concentration of 6-10 vol%;
Step 6, regenerating the lean organic phase
Mixing the lean organic phase with the regenerant according to the volume ratio of (1-3) to 1, carrying out countercurrent regeneration at the temperature of 10-30 ℃ and separating phases to obtain regenerated liquid and regenerated organic phase;
returning the regenerated liquid to the step 5 to be used as a back extraction liquid;
Returning the regenerated organic phase to the step 5 to be used as an organic phase for recycling;
the regenerant is 8-15 vol% sulfuric acid solution;
step 7, precipitating vanadium and preparing vanadium pentoxide product
Uniformly mixing the vanadium-rich liquid and sodium chlorate according to the molar ratio of vanadium to chlorine of (3-6) to 1, and stirring and reacting for 1-3 hours at 50-70 ℃ to obtain the vanadium-rich liquid;
uniformly mixing the vanadium-rich oxide solution and ammonium salt according to the molar ratio of vanadium to nitrogen of 1:2-6, regulating the pH value to 1.8-2.2 by sulfuric acid, stirring for 1-3 h at 75-95 ℃, filtering, and washing to obtain ammonium polyvanadate and vanadium precipitation mother liquor;
deamination is carried out on the ammonium polyvanadate for 1-3 hours at the temperature of 450-550 ℃ to obtain a vanadium pentoxide product;
the vanadium precipitation mother liquor returns to the step 2 and is mixed with the vanadium-containing pickle liquor;
the ammonium salt is one of ammonium chloride, ammonium sulfate and ammonia water.
2. The method for recovering vanadium and iron from vanadium shale according to claim 1, wherein the concentration of the concentrated sulfuric acid is more than or equal to 98wt%.
3. The method for recovering vanadium and iron from vanadium shale according to claim 1, wherein the fluorine-containing leaching aid is one or more of calcium fluoride, sodium fluoride and potassium fluoride.
4. The method for recovering vanadium and iron from vanadium shale according to claim 1, wherein the content of V 2O5 in the vanadium shale is 0.3-3 wt%; the content of Fe 2O3 is 1-10wt%.
5. The method for recovering vanadium and iron from vanadium shale according to claim 1, wherein the phosphorus-containing reducing agent is one or more of sodium hypophosphite, sodium phosphite, hypophosphorous acid and phosphorous acid.
6. The method for recovering vanadium and iron from vanadium shale according to claim 1, wherein the stirring speed is 100-500 r/min.
7. The method for recovering vanadium and iron from vanadium shale according to claim 1, wherein the aerated gas is air or oxygen.
8. The method for recovering vanadium and iron from vanadium shale according to claim 1, wherein the concentration of the sodium sulfate solution is 1-5 wt%.
9. The method for recovering vanadium and iron from vanadium shale according to claim 1, wherein the phosphoric acid extractant is di (2-ethylhexyl) phosphate or mono 2-ethylhexyl phosphate.
10. The method for recovering vanadium and iron from vanadium shale according to claim 1, wherein the phase modifier is sec-octanol or tributyl phosphate.
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