CN115522076B - Method for preparing ammonium metavanadate and vanadium pentoxide from vanadium-containing metallurgical waste residues - Google Patents
Method for preparing ammonium metavanadate and vanadium pentoxide from vanadium-containing metallurgical waste residues Download PDFInfo
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- CN115522076B CN115522076B CN202211331246.8A CN202211331246A CN115522076B CN 115522076 B CN115522076 B CN 115522076B CN 202211331246 A CN202211331246 A CN 202211331246A CN 115522076 B CN115522076 B CN 115522076B
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- vanadium
- sulfuric acid
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- 229910052720 vanadium Inorganic materials 0.000 title claims abstract description 392
- LEONUFNNVUYDNQ-UHFFFAOYSA-N vanadium atom Chemical compound [V] LEONUFNNVUYDNQ-UHFFFAOYSA-N 0.000 title claims abstract description 372
- 239000002699 waste material Substances 0.000 title claims abstract description 71
- 238000000034 method Methods 0.000 title claims abstract description 63
- GNTDGMZSJNCJKK-UHFFFAOYSA-N divanadium pentaoxide Chemical compound O=[V](=O)O[V](=O)=O GNTDGMZSJNCJKK-UHFFFAOYSA-N 0.000 title claims abstract description 54
- UNTBPXHCXVWYOI-UHFFFAOYSA-O azanium;oxido(dioxo)vanadium Chemical compound [NH4+].[O-][V](=O)=O UNTBPXHCXVWYOI-UHFFFAOYSA-O 0.000 title claims abstract description 28
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims abstract description 293
- 238000002386 leaching Methods 0.000 claims abstract description 170
- 239000007788 liquid Substances 0.000 claims abstract description 159
- 238000001556 precipitation Methods 0.000 claims abstract description 100
- 238000003795 desorption Methods 0.000 claims abstract description 96
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims abstract description 94
- 239000011347 resin Substances 0.000 claims abstract description 55
- 229920005989 resin Polymers 0.000 claims abstract description 55
- 239000012452 mother liquor Substances 0.000 claims abstract description 53
- 239000003456 ion exchange resin Substances 0.000 claims abstract description 32
- 229920003303 ion-exchange polymer Polymers 0.000 claims abstract description 32
- NWUYHJFMYQTDRP-UHFFFAOYSA-N 1,2-bis(ethenyl)benzene;1-ethenyl-2-ethylbenzene;styrene Chemical compound C=CC1=CC=CC=C1.CCC1=CC=CC=C1C=C.C=CC1=CC=CC=C1C=C NWUYHJFMYQTDRP-UHFFFAOYSA-N 0.000 claims abstract description 31
- 229920006395 saturated elastomer Polymers 0.000 claims abstract description 29
- 239000011259 mixed solution Substances 0.000 claims abstract description 15
- OYPRJOBELJOOCE-UHFFFAOYSA-N Calcium Chemical compound [Ca] OYPRJOBELJOOCE-UHFFFAOYSA-N 0.000 claims abstract description 12
- 239000011575 calcium Substances 0.000 claims abstract description 12
- 229910052791 calcium Inorganic materials 0.000 claims abstract description 12
- BJVWCKXHSNBHGB-UHFFFAOYSA-L disodium;chloride;hydroxide Chemical compound [OH-].[Na+].[Na+].[Cl-] BJVWCKXHSNBHGB-UHFFFAOYSA-L 0.000 claims abstract description 12
- 239000000243 solution Substances 0.000 claims description 93
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 81
- 239000002893 slag Substances 0.000 claims description 63
- 238000006243 chemical reaction Methods 0.000 claims description 33
- 238000002156 mixing Methods 0.000 claims description 27
- 238000000926 separation method Methods 0.000 claims description 27
- 230000001105 regulatory effect Effects 0.000 claims description 24
- 239000003957 anion exchange resin Substances 0.000 claims description 23
- NLXLAEXVIDQMFP-UHFFFAOYSA-N Ammonia chloride Chemical compound [NH4+].[Cl-] NLXLAEXVIDQMFP-UHFFFAOYSA-N 0.000 claims description 16
- QAOWNCQODCNURD-UHFFFAOYSA-L Sulfate Chemical compound [O-]S([O-])(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-L 0.000 claims description 16
- 239000003513 alkali Substances 0.000 claims description 13
- 238000004090 dissolution Methods 0.000 claims description 12
- BZSXEZOLBIJVQK-UHFFFAOYSA-N 2-methylsulfonylbenzoic acid Chemical compound CS(=O)(=O)C1=CC=CC=C1C(O)=O BZSXEZOLBIJVQK-UHFFFAOYSA-N 0.000 claims description 11
- 238000001354 calcination Methods 0.000 claims description 11
- 239000003795 chemical substances by application Substances 0.000 claims description 10
- QGZKDVFQNNGYKY-UHFFFAOYSA-N Ammonia Chemical compound N QGZKDVFQNNGYKY-UHFFFAOYSA-N 0.000 claims description 8
- VWBLQUSTSLXQON-UHFFFAOYSA-N N.[V+5] Chemical compound N.[V+5] VWBLQUSTSLXQON-UHFFFAOYSA-N 0.000 claims description 8
- FAPWRFPIFSIZLT-UHFFFAOYSA-M Sodium chloride Chemical compound [Na+].[Cl-] FAPWRFPIFSIZLT-UHFFFAOYSA-M 0.000 claims description 8
- 235000019270 ammonium chloride Nutrition 0.000 claims description 8
- 150000003863 ammonium salts Chemical class 0.000 claims description 8
- 238000007654 immersion Methods 0.000 claims description 8
- 239000000463 material Substances 0.000 claims description 7
- 230000001376 precipitating effect Effects 0.000 claims description 7
- NIXOWILDQLNWCW-UHFFFAOYSA-N acrylic acid group Chemical group C(C=C)(=O)O NIXOWILDQLNWCW-UHFFFAOYSA-N 0.000 claims description 6
- 229910021529 ammonia Inorganic materials 0.000 claims description 4
- 239000011780 sodium chloride Substances 0.000 claims description 4
- 239000002253 acid Substances 0.000 abstract description 18
- 238000004064 recycling Methods 0.000 abstract description 5
- 230000007062 hydrolysis Effects 0.000 abstract description 4
- 238000006460 hydrolysis reaction Methods 0.000 abstract description 4
- 238000001179 sorption measurement Methods 0.000 description 87
- 238000005406 washing Methods 0.000 description 39
- 239000007787 solid Substances 0.000 description 32
- 230000008569 process Effects 0.000 description 19
- 238000005342 ion exchange Methods 0.000 description 16
- 238000002791 soaking Methods 0.000 description 15
- VTYYLEPIZMXCLO-UHFFFAOYSA-L Calcium carbonate Chemical compound [Ca+2].[O-]C([O-])=O VTYYLEPIZMXCLO-UHFFFAOYSA-L 0.000 description 12
- 238000001914 filtration Methods 0.000 description 12
- 238000003756 stirring Methods 0.000 description 11
- XEEYBQQBJWHFJM-UHFFFAOYSA-N iron Substances [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 9
- 238000002336 sorption--desorption measurement Methods 0.000 description 9
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 8
- 239000002245 particle Substances 0.000 description 8
- 238000003860 storage Methods 0.000 description 8
- 238000012360 testing method Methods 0.000 description 8
- 238000005070 sampling Methods 0.000 description 7
- 238000000967 suction filtration Methods 0.000 description 7
- PPBRXRYQALVLMV-UHFFFAOYSA-N Styrene Chemical compound C=CC1=CC=CC=C1 PPBRXRYQALVLMV-UHFFFAOYSA-N 0.000 description 6
- 229910000019 calcium carbonate Inorganic materials 0.000 description 6
- 230000000694 effects Effects 0.000 description 6
- 150000002500 ions Chemical class 0.000 description 6
- 229910052742 iron Inorganic materials 0.000 description 6
- 239000011701 zinc Substances 0.000 description 6
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 description 5
- 238000001816 cooling Methods 0.000 description 5
- 238000001035 drying Methods 0.000 description 5
- 239000000706 filtrate Substances 0.000 description 5
- 230000035515 penetration Effects 0.000 description 5
- 238000005086 pumping Methods 0.000 description 5
- 239000000126 substance Substances 0.000 description 5
- 238000012546 transfer Methods 0.000 description 5
- 229910052725 zinc Inorganic materials 0.000 description 5
- LFQSCWFLJHTTHZ-UHFFFAOYSA-N Ethanol Chemical compound CCO LFQSCWFLJHTTHZ-UHFFFAOYSA-N 0.000 description 4
- 239000004677 Nylon Substances 0.000 description 4
- OSGAYBCDTDRGGQ-UHFFFAOYSA-L calcium sulfate Chemical compound [Ca+2].[O-]S([O-])(=O)=O OSGAYBCDTDRGGQ-UHFFFAOYSA-L 0.000 description 4
- 230000000052 comparative effect Effects 0.000 description 4
- 238000011010 flushing procedure Methods 0.000 description 4
- 229910052500 inorganic mineral Inorganic materials 0.000 description 4
- 239000011707 mineral Substances 0.000 description 4
- 239000010413 mother solution Substances 0.000 description 4
- 229920001778 nylon Polymers 0.000 description 4
- 230000032683 aging Effects 0.000 description 3
- 230000008901 benefit Effects 0.000 description 3
- 239000013078 crystal Substances 0.000 description 3
- 239000012153 distilled water Substances 0.000 description 3
- 238000000227 grinding Methods 0.000 description 3
- 238000009776 industrial production Methods 0.000 description 3
- -1 iron ions Chemical class 0.000 description 3
- 229910052751 metal Inorganic materials 0.000 description 3
- 239000000203 mixture Substances 0.000 description 3
- 230000001590 oxidative effect Effects 0.000 description 3
- 238000011084 recovery Methods 0.000 description 3
- SUKJFIGYRHOWBL-UHFFFAOYSA-N sodium hypochlorite Chemical compound [Na+].Cl[O-] SUKJFIGYRHOWBL-UHFFFAOYSA-N 0.000 description 3
- 230000003068 static effect Effects 0.000 description 3
- 229910001456 vanadium ion Inorganic materials 0.000 description 3
- BHPQYMZQTOCNFJ-UHFFFAOYSA-N Calcium cation Chemical compound [Ca+2] BHPQYMZQTOCNFJ-UHFFFAOYSA-N 0.000 description 2
- XUIMIQQOPSSXEZ-UHFFFAOYSA-N Silicon Chemical compound [Si] XUIMIQQOPSSXEZ-UHFFFAOYSA-N 0.000 description 2
- 238000004458 analytical method Methods 0.000 description 2
- 150000001450 anions Chemical class 0.000 description 2
- AXCZMVOFGPJBDE-UHFFFAOYSA-L calcium dihydroxide Chemical compound [OH-].[OH-].[Ca+2] AXCZMVOFGPJBDE-UHFFFAOYSA-L 0.000 description 2
- 239000000920 calcium hydroxide Substances 0.000 description 2
- 229910001861 calcium hydroxide Inorganic materials 0.000 description 2
- 229910001424 calcium ion Inorganic materials 0.000 description 2
- BRPQOXSCLDDYGP-UHFFFAOYSA-N calcium oxide Chemical compound [O-2].[Ca+2] BRPQOXSCLDDYGP-UHFFFAOYSA-N 0.000 description 2
- 239000000292 calcium oxide Substances 0.000 description 2
- ODINCKMPIJJUCX-UHFFFAOYSA-N calcium oxide Inorganic materials [Ca]=O ODINCKMPIJJUCX-UHFFFAOYSA-N 0.000 description 2
- 230000007423 decrease Effects 0.000 description 2
- 238000009792 diffusion process Methods 0.000 description 2
- 238000002149 energy-dispersive X-ray emission spectroscopy Methods 0.000 description 2
- 238000000605 extraction Methods 0.000 description 2
- 239000006260 foam Substances 0.000 description 2
- 230000003301 hydrolyzing effect Effects 0.000 description 2
- JQJCSZOEVBFDKO-UHFFFAOYSA-N lead zinc Chemical compound [Zn].[Pb] JQJCSZOEVBFDKO-UHFFFAOYSA-N 0.000 description 2
- 238000004519 manufacturing process Methods 0.000 description 2
- 230000010198 maturation time Effects 0.000 description 2
- 230000007246 mechanism Effects 0.000 description 2
- 239000002184 metal Substances 0.000 description 2
- FWFGVMYFCODZRD-UHFFFAOYSA-N oxidanium;hydrogen sulfate Chemical compound O.OS(O)(=O)=O FWFGVMYFCODZRD-UHFFFAOYSA-N 0.000 description 2
- 230000003647 oxidation Effects 0.000 description 2
- 238000007254 oxidation reaction Methods 0.000 description 2
- 239000000047 product Substances 0.000 description 2
- 238000000746 purification Methods 0.000 description 2
- 239000002994 raw material Substances 0.000 description 2
- 238000001878 scanning electron micrograph Methods 0.000 description 2
- 238000012216 screening Methods 0.000 description 2
- 229910052710 silicon Inorganic materials 0.000 description 2
- 239000010703 silicon Substances 0.000 description 2
- 239000008399 tap water Substances 0.000 description 2
- 235000020679 tap water Nutrition 0.000 description 2
- QGZKDVFQNNGYKY-UHFFFAOYSA-O Ammonium Chemical compound [NH4+] QGZKDVFQNNGYKY-UHFFFAOYSA-O 0.000 description 1
- 229910004283 SiO 4 Inorganic materials 0.000 description 1
- 239000003929 acidic solution Substances 0.000 description 1
- 230000002378 acidificating effect Effects 0.000 description 1
- 230000006978 adaptation Effects 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 230000015572 biosynthetic process Effects 0.000 description 1
- 239000003054 catalyst Substances 0.000 description 1
- 230000008859 change Effects 0.000 description 1
- 238000009388 chemical precipitation Methods 0.000 description 1
- 239000012295 chemical reaction liquid Substances 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000010494 dissociation reaction Methods 0.000 description 1
- 230000005593 dissociations Effects 0.000 description 1
- 238000005265 energy consumption Methods 0.000 description 1
- 238000005516 engineering process Methods 0.000 description 1
- 238000001704 evaporation Methods 0.000 description 1
- 230000002349 favourable effect Effects 0.000 description 1
- 239000012527 feed solution Substances 0.000 description 1
- IMBKASBLAKCLEM-UHFFFAOYSA-L ferrous ammonium sulfate (anhydrous) Chemical compound [NH4+].[NH4+].[Fe+2].[O-]S([O-])(=O)=O.[O-]S([O-])(=O)=O IMBKASBLAKCLEM-UHFFFAOYSA-L 0.000 description 1
- 238000011049 filling Methods 0.000 description 1
- 229910052745 lead Inorganic materials 0.000 description 1
- 229910044991 metal oxide Inorganic materials 0.000 description 1
- 150000004706 metal oxides Chemical class 0.000 description 1
- 150000002739 metals Chemical class 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 238000000643 oven drying Methods 0.000 description 1
- 238000006400 oxidative hydrolysis reaction Methods 0.000 description 1
- 238000010979 pH adjustment Methods 0.000 description 1
- 238000012856 packing Methods 0.000 description 1
- 230000000149 penetrating effect Effects 0.000 description 1
- 238000006116 polymerization reaction Methods 0.000 description 1
- 239000000843 powder Substances 0.000 description 1
- 230000002035 prolonged effect Effects 0.000 description 1
- 230000035484 reaction time Effects 0.000 description 1
- KKCBUQHMOMHUOY-UHFFFAOYSA-N sodium oxide Chemical compound [O-2].[Na+].[Na+] KKCBUQHMOMHUOY-UHFFFAOYSA-N 0.000 description 1
- 229910001948 sodium oxide Inorganic materials 0.000 description 1
- 238000007614 solvation Methods 0.000 description 1
- 239000002904 solvent Substances 0.000 description 1
- 230000008961 swelling Effects 0.000 description 1
- 230000002195 synergetic effect Effects 0.000 description 1
- 238000004448 titration Methods 0.000 description 1
- DIMMBYOINZRKMD-UHFFFAOYSA-N vanadium(5+) Chemical compound [V+5] DIMMBYOINZRKMD-UHFFFAOYSA-N 0.000 description 1
- 238000004846 x-ray emission Methods 0.000 description 1
Classifications
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B34/00—Obtaining refractory metals
- C22B34/20—Obtaining niobium, tantalum or vanadium
- C22B34/22—Obtaining vanadium
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01G—COMPOUNDS CONTAINING METALS NOT COVERED BY SUBCLASSES C01D OR C01F
- C01G31/00—Compounds of vanadium
- C01G31/003—Preparation involving a liquid-liquid extraction, an adsorption or an ion-exchange
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01G—COMPOUNDS CONTAINING METALS NOT COVERED BY SUBCLASSES C01D OR C01F
- C01G31/00—Compounds of vanadium
- C01G31/02—Oxides
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
- C22B3/22—Treatment or purification of solutions, e.g. obtained by leaching by physical processes, e.g. by filtration, by magnetic means, or by thermal decomposition
- C22B3/24—Treatment or purification of solutions, e.g. obtained by leaching by physical processes, e.g. by filtration, by magnetic means, or by thermal decomposition by adsorption on solid substances, e.g. by extraction with solid resins
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
- C22B3/42—Treatment or purification of solutions, e.g. obtained by leaching by ion-exchange extraction
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/001—Dry processes
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/006—Wet processes
- C22B7/007—Wet processes by acid leaching
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/006—Wet processes
- C22B7/008—Wet processes by an alkaline or ammoniacal leaching
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/04—Working-up slag
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
Abstract
The invention provides a method for preparing ammonium metavanadate and vanadium pentoxide by utilizing vanadium-containing metallurgical waste residues, and relates to the technical field of vanadium resource recycling. The method adopts a sulfuric acid curing-water leaching mode to fully dissolve vanadium in vanadium-containing metallurgical waste residues, the acid consumption is low, then an alkaline calcium source is utilized to remove sulfate radicals, the yield of vanadium and the purity of products are improved, then red vanadium and precipitation mother liquor are obtained through hydrolysis and precipitation of vanadium, the precipitation mother liquor is desorbed by a sulfuric acid type ion exchange resin after being recovered by a sodium hydroxide-sodium chloride mixed solution, primary vanadium-lean desorption liquid, vanadium-rich desorption liquid and secondary vanadium-lean desorption liquid are sequentially obtained, the vanadium-rich desorption liquid is used for dissolving red vanadium, the pH values of the primary vanadium-lean desorption liquid and the secondary vanadium-lean desorption liquid are adjusted to 13-14 and then are used for desorption of saturated resin of the next batch, and the full recycling of vanadium resources in the precipitation mother liquor is realized, so that the total yield of vanadium is greatly improved.
Description
Technical Field
The invention relates to the technical field of vanadium resource recycling, in particular to a method for preparing ammonium metavanadate and vanadium pentoxide by utilizing vanadium-containing metallurgical waste residues.
Background
At present, chemical precipitation technology is mainly adopted for treating vanadium in vanadium leaching solution in industrial production, and comprises ammonium salt vanadium precipitation, hydrolysis vanadium precipitation and the like. For example, chinese patent CN102121068A discloses a process for preparing vanadium pentoxide comprising the steps of: (1) Adjusting the pH value of the vanadium-containing acidic solution to 4.0-4.5, and carrying out precipitation and filtration; (2) Oxidizing and alkaline leaching the filter residue obtained in the step (1) to obtain alkaline leaching liquid and residue; (3) Regulating the pH value of alkaline leaching solution to 1.8-2.2, adding ammonium salt, and carrying out vanadium precipitation reaction to prepare ammonium vanadate vanadium slag; (4) Calcining ammonium vanadate vanadium slag to obtain vanadium pentoxide with purity of 99.91-99.95%, and obtaining vanadium pentoxide with purity of 99.91-99.95%. However, the filtrate obtained in the step (1) of the prior art has residual vanadium, and the concentration of vanadium in the filtrate is low, so that the vanadium in the filtrate is difficult to recover by a precipitation method.
Chinese patent CN110438336a discloses a method for extracting vanadium pentoxide from vanadium-containing lead zinc ore, comprising the steps of: drying and grinding vanadium lead zinc ore, adding sulfuric acid solution for acidic leaching, carrying out solid-liquid separation after leaching to obtain filtrate and filter residue, adding sodium chlorate into the filtrate for oxidation, carrying out ion exchange on leaching solution subjected to oxidation treatment, adsorbing vanadium ions in resin after ion exchange is completed, and obtaining solution with lower vanadium ion concentration after ion exchange; desorbing the ion-exchanged resin by using sodium hydroxide solution, evaporating and concentrating the obtained vanadium-containing desorption solution, adding ammonium salt, adjusting the pH value to be 2.0-2.2 so as to precipitate vanadium from the ammonium salt, and calcining the obtained ammonium metavanadate solid to obtain vanadium pentoxide. However, the ion exchange in the prior art is only suitable for treating low-concentration vanadium-containing liquid, the concentration is too high, the adsorption saturation is faster, the industrial production and the application are not facilitated, and the vanadium in the obtained adsorbed liquid is not reasonably recycled, so that the total vanadium yield in the prior art is not high enough (about 80%).
Disclosure of Invention
In view of the above, the present invention aims to provide a method for preparing ammonium metavanadate and vanadium pentoxide by utilizing vanadium-containing metallurgical waste residues, wherein vanadium in solutions with different vanadium concentrations is recovered step by step, vanadium in vanadium-containing sulfuric acid leaching solution (high-concentration vanadium) is precipitated by oxidative hydrolysis, and vanadium in precipitation mother liquor (low-concentration vanadium) is adsorbed and recovered by sulfuric acid type ion exchange resin, so that the total vanadium yield (more than 90%) is high.
In order to achieve the above object, the present invention provides the following technical solutions:
the invention provides a method for preparing ammonium metavanadate from vanadium-containing metallurgical waste residues, which comprises the following steps:
(1) Mixing vanadium-containing metallurgical waste residues with concentrated sulfuric acid, curing, mixing the obtained cured material with water, and leaching with water to obtain vanadium-containing sulfuric acid leaching solution;
(2) Removing sulfate radical in the vanadium-containing sulfuric acid leaching solution by using an alkaline calcium source to obtain sulfate radical-removing vanadium-containing sulfuric acid leaching solution; mixing the sulfate radical-removing vanadium-containing sulfuric acid leaching solution with sodium hydroxide and sodium chlorate to perform oxidation-vanadium precipitation reaction, and carrying out solid-liquid separation to obtain red vanadium and precipitation mother liquor respectively; the pH value of the oxidation-vanadium precipitation reaction is 2-2.4;
(3) Absorbing the precipitation mother liquor to saturation by utilizing sulfuric acid type ion exchange resin, and then desorbing by utilizing an alkaline desorbing agent to sequentially obtain primary vanadium-lean desorption liquid, vanadium-rich desorption liquid and secondary vanadium-lean desorption liquid; adjusting the pH values of the primary vanadium-lean desorption liquid and the secondary vanadium-lean desorption liquid to 13-14 by using sodium hydroxide, and then desorbing the saturated resin in the next batch; the volume ratio of the primary vanadium-lean desorption liquid to the secondary vanadium-lean desorption liquid is 1-1.5: 1.5 to 3.5:3.5 to 6; the alkaline desorbent is a sodium hydroxide-sodium chloride mixed solution;
(4) Alkali dissolution is carried out on the red vanadium by utilizing the vanadium-rich desorption liquid to obtain vanadium-rich liquid;
(5) And regulating the pH value of the vanadium-rich liquid to 8-9 by utilizing sulfuric acid, mixing with ammonium chloride, and carrying out ammonium salt vanadium precipitation to obtain ammonium metavanadate.
Preferably, in the step (1), the mass ratio of the vanadium-containing metallurgical waste residue to the concentrated sulfuric acid is 1:0.6 to 0.7;
the curing time is 2-5 h.
Preferably, in the step (1), the ratio of the mass of the vanadium-containing metallurgical slag to the volume of water is 1g: 3-6 mL;
the water immersion time is 1-2.5 h.
Preferably, in the step (2), the mass ratio of the vanadium-containing sulfuric acid leaching solution to sodium chlorate is 1:0.02 to 0.1, wherein the mass of the vanadium-containing sulfuric acid leaching solution is expressed as V 2 O 5 Counting;
the temperature of the oxidation-vanadium precipitation reaction is 90-100 ℃ and the time is 1.5-2 h.
Preferably, in the step (3), the sulfuric acid type ion exchange resin comprises at least one of sulfuric acid type D231-YT macroporous strongly basic anion exchange resin, sulfuric acid type D301 macroporous weakly basic styrenic anion exchange resin, and sulfuric acid type D314 macroporous weakly basic acrylic anion exchange resin.
Preferably, in the step (3), the concentration of sodium hydroxide in the sodium hydroxide-sodium chloride mixed solution is 5-7.5%, and the concentration of sodium chloride is 8-10%.
Preferably, in the step (4), the ratio of the volume of the vanadium-rich desorption liquid to the dry weight of red vanadium is 6-8 mL:1g;
the alkali dissolution temperature is 90-100 ℃, the time is 2-3 h, and the pH value is 13-14.
Preferably, in the step (5), the ammonia addition coefficient of the ammonium chloride is 3-4;
the temperature of the ammonium salt for precipitating vanadium is 90-100 ℃ and the time is 1-2 h.
The invention provides a method for preparing vanadium pentoxide by utilizing vanadium-containing metallurgical waste residues, which comprises the following steps:
calcining the ammonium metavanadate prepared by the method of the technical scheme to obtain vanadium pentoxide.
Preferably, the calcination temperature is 500-600 ℃ and the time is 2-3 h.
The invention provides a method for preparing ammonium metavanadate from vanadium-containing metallurgical waste residues, which comprises the following steps: (1) Mixing vanadium-containing metallurgical waste residues with concentrated sulfuric acid, curing, mixing the obtained cured material with water, and leaching with water to obtain vanadium-containing sulfuric acid leaching solution; (2) Removing sulfate radical in the vanadium-containing sulfuric acid leaching solution by using an alkaline calcium source to obtain sulfate radical-removing vanadium-containing sulfuric acid leaching solution; mixing the sulfate radical-removing vanadium-containing sulfuric acid leaching solution with sodium hydroxide and sodium chlorate to perform oxidation-vanadium precipitation reaction, and carrying out solid-liquid separation to obtain red vanadium and precipitation mother liquor respectively; the pH value of the oxidation-vanadium precipitation reaction is 2-2.4; (3) Absorbing the precipitation mother liquor to saturation by utilizing sulfuric acid type ion exchange resin, and then desorbing by utilizing an alkaline desorbing agent to sequentially obtain primary vanadium-lean desorption liquid, vanadium-rich desorption liquid and secondary vanadium-lean desorption liquid; adjusting the pH values of the primary vanadium-lean desorption liquid and the secondary vanadium-lean desorption liquid to 13-14 by using sodium hydroxide, and then desorbing the saturated resin in the next batch; the volume ratio of the primary vanadium-lean desorption liquid to the secondary vanadium-lean desorption liquid is 1-1.5: 1.5 to 3.5:3.5 to 6; the alkaline desorbent is a sodium hydroxide-sodium chloride mixed solution; (4) Alkali dissolution is carried out on the red vanadium by utilizing the vanadium-rich desorption liquid to obtain vanadium-rich liquid; (5) And regulating the pH value of the vanadium-rich liquid to 8-9 by utilizing sulfuric acid, mixing with ammonium chloride, and carrying out ammonium salt vanadium precipitation to obtain ammonium metavanadate. The method adopts a sulfuric acid curing-water leaching mode to fully dissolve vanadium in the vanadium-containing metallurgical waste residue, and then utilizes an alkaline calcium source to remove sulfate radicals, so that the influence of the excessive sulfate radical concentration on the vanadium precipitation rate is avoided, and the purity of ammonium metavanadate can be improved; and then recovering vanadium in solutions with different concentrations step by step, specifically, precipitating vanadium in vanadium-containing sulfuric acid leaching solution of high-concentration vanadium (namely, oxidizing low-valence vanadium (+ 3, +4) in a system into high-valence vanadium (+ 5) by utilizing sodium chlorate and precipitating vanadium), so as to form red vanadium with low vanadium grade (40-50%), recovering a precipitation mother solution of the low-concentration vanadium by utilizing sulfuric acid ion exchange resin, then desorbing by utilizing sodium hydroxide-sodium chloride mixed solution, recycling vanadium-rich desorption solution obtained by desorption for dissolving the red vanadium, regulating the pH value of other desorption solutions (primary vanadium-lean desorption solution and secondary vanadium-lean desorption solution) to 13-14 and then using the desorption of saturated resin of the next batch, thereby realizing full recycling of vanadium resources in the precipitation mother solution, greatly improving the total yield of vanadium, and simultaneously providing a new way for improving the hydrolysis vanadium precipitation process and having good economic benefits. Compared with the traditional direct acid leaching process, the method adopts the sulfuric acid curing-water leaching process, can obviously reduce the acid consumption and improve the vanadium leaching rate, and has good economic benefit. The method provided by the invention aims at the synergistic purification of vanadium solutions with different concentrations, and has the advantages of stepwise enrichment, large treatment capacity, strong adaptability, total vanadium yield of more than 90% and high total vanadium yield. Moreover, the method provided by the invention is simple to operate, low in cost and suitable for industrial production.
The invention also provides a method for preparing vanadium pentoxide by utilizing the vanadium-containing metallurgical waste residue, which comprises the step of calcining the ammonium metavanadate prepared by the method in the technical scheme to obtain the vanadium pentoxide. As shown in the test results of the examples, the method uses V in the vanadium-containing metallurgical waste slag 2 O 5 The total yield of vanadium in the method provided by the invention is more than 90 percent.
Drawings
FIG. 1 is a diagram of a column type dynamic adsorption-desorption apparatus in which a 1-feed liquid storage tank, a 2-transfer pump, a 3-constant pressure tank, a 4-vent, a 5-nylon cushion layer, a 6-adsorption column, 7-resin, an 8-regulating valve, and a 9-automatic collection sampler are provided;
FIG. 2 shows the pH versus sulfuric acid type D314 resin versus V 2 O 5 An influence map of the saturated adsorption amount of (2);
FIG. 3 shows adsorption temperature vs. V 2 O 5 The influence of the saturated adsorption amount of (2);
FIG. 4 is a graph showing adsorption time versus V 2 O 5 The influence of the saturated adsorption amount of (2);
FIG. 5 is a column dynamic adsorption V 2 O 5 A graph;
FIG. 6 is a column dynamic desorption V 2 O 5 A graph;
FIG. 7 is a flow chart of a process for preparing vanadium pentoxide;
FIG. 8 shows the sulfuric acid ripening-water leaching parameters versus vanadium (in V 2 O 5 Metering) the influence of leaching rate is shown in figure 8, wherein (a) is the influence of concentrated sulfuric acid consumption on vanadium leaching, (b) is the influence of curing time on vanadium leaching, (c) is the influence of vanadium-containing metallurgical waste residue granularity on vanadium leaching, (d) is the influence of water leaching liquid-solid ratio on vanadium leaching, and (e) is the influence of water leaching time on vanadium leaching;
FIG. 9 is an SEM image of the vanadium-containing metallurgical slag, the resulting slaked slag and sulfuric acid-aged water-immersed slag used in example 2, and the acid-immersed slag produced in comparative example 1, wherein (a) is the vanadium-containing metallurgical slag, (b) is the acid-immersed slag, (c) is the aged slag, and (d) is the sulfuric acid-aged water-immersed slag;
FIG. 10 is an EDS analysis chart of the vanadium-containing metallurgical slag and the sulfuric acid-water leaching slag obtained in example 2, wherein (a) is the vanadium-containing metallurgical slag and (b) is the sulfuric acid-water leaching slag.
Detailed Description
The invention provides a method for preparing ammonium metavanadate from vanadium-containing metallurgical waste residues, which comprises the following steps:
(1) Mixing vanadium-containing metallurgical waste residues with concentrated sulfuric acid, curing, mixing the obtained cured material with water, and leaching with water to obtain vanadium-containing sulfuric acid leaching solution;
(2) Removing sulfate radical in the vanadium-containing sulfuric acid leaching solution by using an alkaline calcium source to obtain sulfate radical-removing vanadium-containing sulfuric acid leaching solution; mixing the sulfate radical-removing vanadium-containing sulfuric acid leaching solution with sodium hydroxide and sodium chlorate to perform oxidation-vanadium precipitation reaction, and carrying out solid-liquid separation to obtain red vanadium and precipitation mother liquor respectively; the pH value of the oxidation-vanadium precipitation reaction is 2-2.4;
(3) Absorbing the precipitation mother liquor to saturation by utilizing sulfuric acid type ion exchange resin, and then desorbing by utilizing an alkaline desorbing agent to sequentially obtain primary vanadium-lean desorption liquid, vanadium-rich desorption liquid and secondary vanadium-lean desorption liquid; adjusting the pH values of the primary vanadium-lean desorption liquid and the secondary vanadium-lean desorption liquid to 13-14 by using sodium hydroxide, and then desorbing the saturated resin in the next batch; the volume ratio of the primary vanadium-lean desorption liquid to the secondary vanadium-lean desorption liquid is 1-1.5: 1.5 to 3.5:3.5 to 6; the alkaline desorbent is a sodium hydroxide-sodium chloride mixed solution;
(4) Alkali dissolution is carried out on the red vanadium by utilizing the vanadium-rich desorption liquid to obtain vanadium-rich liquid;
(5) And regulating the pH value of the vanadium-rich liquid to 8-9 by utilizing sulfuric acid, mixing with ammonium chloride, and carrying out ammonium salt vanadium precipitation to obtain ammonium metavanadate.
In the present invention, all raw material components are commercially available products well known to those skilled in the art unless specified otherwise.
The invention mixes the vanadium-containing metallurgical waste residue with concentrated sulfuric acid for curing, mixes the obtained curing material with water, and carries out water leaching to obtain vanadium-containing sulfuric acid leaching solution.
In the invention, the mass of the vanadium-containing metallurgical waste residue with the granularity of-0.6 mm is preferably more than 90% of the total mass of the vanadium-containing metallurgical waste residue, more preferably-0.18 mm is preferably more than 90% of the total mass of the vanadium-containing metallurgical waste residue, and further preferably-0.125 mm is preferably more than 90% of the total mass of the vanadium-containing metallurgical waste residue. In the present invention, the concentration of the concentrated sulfuric acid is preferably 98wt%. In the invention, the mass ratio of the vanadium-containing metallurgical waste residue to the concentrated sulfuric acid is preferably 1:0.6 to 0.7, more preferably 1:0.7. the invention is not particularly limited to the vanadium-containing metallurgical slag, and the vanadium-containing metallurgical slag well known to those skilled in the art can be used. In the specific embodiment of the invention, the main chemical components of the vanadium-containing metallurgical waste residues are shown in the following table 1:
TABLE 1 main chemical Components of vanadium-containing Metallurgical slag
In the invention, the curing temperature is the temperature at which concentrated sulfuric acid is added into vanadium-containing metallurgical waste residue at room temperature to release heat in the curing process; the curing time is preferably 2 to 5 hours, more preferably 3 to 4 hours. The method utilizes the heat generated by the reaction per se in the curing process, reduces the consumption of energy sources, and the water immersion is carried out at normal temperature and normal pressure, so that the method has low requirements on equipment, simple process and high vanadium extraction rate; in addition, compared with dilute sulfuric acid, the reaction force of the concentrated sulfuric acid curing reaction is large, the damage to the structure of the vanadium-containing metallurgical waste residue is relatively sufficient, and the strong oxidizing property of the concentrated sulfuric acid oxidizes low-valence vanadium into high-valence vanadium, so that the vanadium is easier to leach.
In the invention, the ratio of the volume of water to the mass of the vanadium-containing metallurgical waste residue is preferably 3-6 mL:1g (i.e., a liquid to solid ratio of 3 to 6:1), more preferably 1g: 4-5 mL.
In the present invention, the temperature of the water immersion is preferably 20 to 30 ℃, more preferably room temperature, and the time of the water immersion is preferably 1 to 2.5 hours, more preferably 1.5 to 2 hours.
After the water leaching, the invention preferably further comprises the steps of carrying out solid-liquid separation on the obtained water leaching system to obtain a liquid component and a solid component, washing the solid component with water, and combining the obtained water washing liquid with the liquid component to obtain the vanadium-containing sulfuric acid leaching liquid. The solid-liquid separation method is not particularly limited, and may be any solid-liquid separation method known to those skilled in the art, such as filtration, suction filtration, or centrifugal separation. The water consumption of the water washing method is not particularly limited, the pH value of the vanadium-containing sulfuric acid leaching solution can be 0.2-1, and the pH value of the vanadium-containing sulfuric acid leaching solution is more preferably 0.3-0.8; in V form 2 O 5 The concentration of vanadium in the vanadium-containing sulfuric acid leaching solution is 15-20 g/L, more preferably 16-18 g/L.
After vanadium-containing sulfuric acid leaching solution is obtained, sulfate radical in the vanadium-containing sulfuric acid leaching solution is removed by utilizing an alkaline calcium source, so that sulfate radical-removing vanadium-containing sulfuric acid leaching solution is obtained; mixing the sulfate radical-removing vanadium-containing sulfuric acid leaching solution with sodium hydroxide and sodium chlorate to perform oxidation-vanadium precipitation reaction, and carrying out solid-liquid separation to obtain red vanadium and precipitation mother liquor respectively; the pH value of the oxidation-vanadium precipitation reaction is 2-2.4.
In the present invention, the removal of sulfate groups from the vanadium-containing sulfuric acid leach solution using an alkaline calcium source is preferably: and (3) regulating the pH value of the vanadium-containing sulfuric acid leaching solution to 1.5-1.8 by utilizing an alkaline calcium source, and performing solid-liquid separation (marked as second solid-liquid separation) to obtain a liquid component, namely the sulfate radical-removing vanadium-containing sulfuric acid leaching solution. In the present invention, the alkaline calcium source preferably includes at least one of calcium carbonate, calcium oxide, and calcium hydroxide, more preferably calcium carbonate, calcium oxide, or calcium hydroxide; the calcium carbonate is preferably heavy calcium carbonate; the amount of the alkaline calcium source used in the present invention is not particularly limited, and the pH of the mixed solution may be adjusted to 1.5 to 1.8 (more preferably 1.6 to 1.7). In the invention, in the sulfate radical removal process, an alkaline calcium source and a vanadium-containing sulfuric acid leaching solution H in (1) + The reaction is carried out to generate calcium ions, the calcium ions react with sulfate radicals to generate calcium sulfate, and the calcium sulfate is removed through solid-liquid separation (marked as third solid-liquid separation). The third solid-liquid separation method is not particularly limited, and any solid-liquid separation method known to those skilled in the art may be used, such as filtration, suction filtration, or centrifugal separation.
In the present invention, the mixing is preferably: mixing the sulfate radical-removing vanadium-containing sulfuric acid leaching solution with sodium chlorate and then mixing with sodium hydroxide. In the present invention, the mixing is preferably stirring and mixing, and the speed and time of the stirring and mixing are not particularly limited, and the raw materials may be uniformly mixed. In the present invention, the pH value of the oxidation-vanadium precipitation reaction is more preferably 2.1 to 2.3, still more preferably 2.2; when the pH value is above 2.4, iron ions are precipitated in the subsequent oxidation-vanadium precipitation reaction process, so that the purity of the ammonium metavanadate product is affected.
In the invention, the mass ratio of the vanadium-containing sulfuric acid leaching solution to sodium chlorate is preferably 1:0.02 to 0.1, more preferably 1:0.04 to 0.08, more preferably 1:0.05 to 0.06; the quality of the sulfate radical removal vanadium-containing sulfuric acid leaching solution is expressed as V 2 O 5 And (5) counting.
In the present invention, the temperature of the oxidation-vanadium precipitation reaction is preferably 90 to 100 ℃, more preferably 92 to 98 ℃, and even more preferably 94 to 95 ℃; the time of the oxidation-vanadium precipitation reaction is preferably 1.5 to 2 hours, more preferably 1.8 to 2 hours. In the invention, in the oxidation-vanadium precipitation reaction process, low-valence vanadium in sulfate radical-removal vanadium-containing sulfuric acid leaching solution is oxidized into pentavalent vanadium, the pentavalent vanadium is precipitated to form red vanadium, and vanadium in the precipitation mother liquor mainly adopts H 2 V 10 O 28 4- In the form of ions.
The mode of the second solid-liquid separation is not particularly limited, and any solid-liquid separation mode known to those skilled in the art may be used, such as filtration, suction filtration or centrifugal separation. After the second solid-liquid separation, the present invention preferably further comprises washing the obtained solid component with water, and subjecting the obtained washing liquid to the second solid-liquid separation to obtain a liquidThe components are combined and then the pH value (namely the pH value of the precipitation mother liquor) is regulated to 2.2-2.4, so as to obtain the precipitation mother liquor. The water consumption of the water washing is not particularly limited, and the method uses vanadium (V in the precipitation mother liquor 2 O 5 Based on the total weight of the composition) is 2-3 g/L. In the present invention, the adjustment of the pH is preferably performed using sodium hydroxide. In the present invention, the pH of the precipitation mother liquor is more preferably 2.2 to 2.3.
In the present invention, the vanadium grade of the red vanadium is preferably 40 to 50%, more preferably 45 to 50%.
After obtaining a precipitation mother liquor, the invention utilizes sulfuric acid type ion exchange resin to adsorb the precipitation mother liquor to saturation, then utilizes an alkaline desorber to desorb, and sequentially obtains a primary vanadium-lean desorption solution, a vanadium-rich desorption solution and a secondary vanadium-lean desorption solution; adjusting the pH values of the primary vanadium-lean desorption liquid and the secondary vanadium-lean desorption liquid to 13-14 by using sodium hydroxide, and then desorbing the saturated resin in the next batch; the volume ratio of the primary vanadium-lean desorption liquid to the secondary vanadium-lean desorption liquid is 1-1.5: 1.5 to 3.5:3.5 to 6; the alkaline desorbent is a sodium hydroxide-sodium chloride mixed solution.
In the present invention, the sulfuric acid type ion exchange resin is preferably obtained by pretreatment of a macroporous basic anion exchange resin, and the pretreatment preferably comprises: sequentially performing water soaking, primary water washing, alcohol soaking, secondary water washing, hydrochloric acid solution soaking, tertiary water washing, sodium hydroxide solution soaking, quaternary water washing and sulfuric acid solution soaking. In the present invention, the macroporous basic anion exchange resin preferably includes at least one of D231-YT macroporous strongly basic anion exchange resin, D301 macroporous weakly basic styrene-based anion exchange resin, and D314 macroporous weakly basic acrylic-based anion exchange resin, more preferably D231-YT macroporous strongly basic anion exchange resin, D301 macroporous weakly basic styrene-based anion exchange resin, or D314 macroporous weakly basic acrylic-based anion exchange resin, and still more preferably D314 macroporous weakly basic acrylic-based anion exchange resin. In the present invention, the temperature of the pretreatment is preferably 20 to 30 ℃, more preferably room temperature. In the present invention, the time of the water soaking is preferably 12 to 36 hours, more preferably 24 hours. The number of times of the one-time washing is not particularly limited, and the water washing is carried out until the macroporous alkaline anion exchange resin has no peculiar smell and the washing liquid has no foam. In the present invention, the time of the alcohol soaking is preferably 6 to 18 hours, more preferably 12 hours. In the present invention, the concentrations of the hydrochloric acid solution and the sodium hydroxide solution are independently preferably 0.2 to 0.8mol/L, more preferably 0.5mol/L; the time of the hydrochloric acid solution soaking and the sodium hydroxide solution soaking is independently preferably 6 to 18 hours, more preferably 12 hours. The number of times of the secondary washing, the tertiary washing and the quaternary washing is not particularly limited, and the washing may be performed until the pH of the washing liquid is 6 to 8 (more preferably 7). In the present invention, the concentration of the sulfuric acid solution is preferably 25 to 35g/L, more preferably 30g/L; the time of the sulfuric acid soaking is preferably 12 to 36 hours, more preferably 24 hours.
In the present invention, the adsorption is preferably a dynamic adsorption by a sulfuric acid type ion exchange resin column, and the packing volume fraction of the sulfuric acid type ion exchange resin in the sulfuric acid type ion exchange resin column is preferably 75 to 85%, more preferably 80%. In a specific embodiment of the present invention, the adsorption is preferably performed using a column type dynamic adsorption-desorption apparatus as shown in fig. 1. In the invention, according to the material flow sequence, the column type dynamic adsorption-desorption device preferably comprises a material liquid storage tank 1, a delivery pump 2, a constant pressure tank 3, an adsorption column 6, a regulating valve 8 and an automatic collection sampler 9 which are sequentially communicated, wherein an exhaust port 4 is preferably arranged on a pipeline communicated between the outlet of the constant pressure tank 3 and the inlet of the adsorption column 6, a nylon cushion layer 5 is arranged in the adsorption column 6, and the surface of the nylon cushion layer 5 is filled with resin 7. The following describes, with reference to fig. 1, a specific process of dynamic adsorption using the column type dynamic adsorption-desorption apparatus shown in fig. 1: placing the precipitation mother liquor in a feed liquid storage tank 1, pumping the precipitation mother liquor into a constant pressure tank 3 by a delivery pump 2, allowing the precipitation mother liquor to flow out from the bottom opening of the constant pressure tank 3, allowing the precipitation mother liquor to enter from the bottom of an adsorption column 6, allowing the precipitation mother liquor to flow through a resin 7 from bottom to top, allowing the precipitation mother liquor to flow out from the top of the adsorption column 6 to an automatic collection sampler 9 for sampling and collection, regulating the outflow flow by a regulating valve 8, and testing vanadium (in V) in the adsorption effluent every 1 column volume (BV) 2 O 5 Meter) concentrationWhen vanadium (in V 2 O 5 Calculated by V) and the concentration of vanadium (expressed as V) in the precipitation mother liquor 2 O 5 Meter), continuing to adsorb for 2-3 BV after the concentration is equal, wherein the sulfuric acid type ion exchange resin is adsorbed and saturated, and then draining and washing the column inner liquid in the sulfuric acid type ion exchange resin column to obtain a saturated resin column and a column inner residual liquid respectively; the column residue is combined with the non-adsorbed precipitation mother liquor. In the present invention, the flow rate of the precipitation mother liquor into the sulfuric acid type ion exchange resin column (i.e., the inlet flow rate) is preferably 0.1 to 0.2mL/min, more preferably 0.15 to 0.18mL/min. In the present invention, the flow rate of the adsorption effluent (i.e., effluent flow rate) is preferably 0.9 to 1.1BV/h, more preferably 1BV/h. In the present invention, the water amount of the water washing is preferably 1 to 2BV. In the present invention, the mechanism of the adsorption is shown as formula (1), wherein R represents a macroporous basic anion exchange resin.
2RSO 4 2- +H 2 V 10 O 28 4- ＝R 2 H 2 V 10 O 28 4- +2SO 4 2- Formula (1).
In the invention, the penetration point of the sulfuric acid type ion exchange resin column is preferably 8-12 BV, more preferably 10BV; the penetrating bed volume of the sulfuric acid type ion exchange resin column is preferably 45-55 BV, more preferably 50BV; the saturated bed volume of the sulfuric acid type ion exchange resin column is preferably 90 to 100BV, more preferably 95 to 100BV.
In the present invention, the concentration of sodium hydroxide in the sodium hydroxide-sodium chloride mixed solution is preferably 5 to 7.5%, more preferably 5 to 7%, still more preferably 5 to 6%; the concentration of sodium oxide in the sodium hydroxide-sodium chloride mixed solution is preferably 8 to 10%, more preferably 8.5 to 10%, and even more preferably 9 to 10%; the solvent in the sodium hydroxide-sodium chloride mixed solution is water. In the present invention, the alkaline desorbent is preferably used in an amount of 6 to 11BV, more preferably 8 to 10BV. In the invention, the volume ratio of the primary vanadium-lean desorption liquid to the secondary vanadium-lean desorption liquid is 1-1.5: 1.5 to 3.5: 3.5-6, specifically, the amount of the primary vanadium-lean desorption liquid is preferably 1-1.5 BV, the amount of the vanadium-rich desorption liquid is preferably 1.5-3.5 BV, and the amount of the secondary vanadium-lean desorption liquid is preferably 3.5-6 BV.
In the invention, the desorption is preferably carried out by using a column type dynamic adsorption-desorption device shown in figure 1, specifically, the alkaline desorber is placed in a feed liquid storage tank 1, pumped into a constant pressure tank 3 by a delivery pump 2, the alkaline desorber flows out from the bottom opening of the constant pressure tank 3, enters from the bottom of an adsorption column 6, flows through resin positioned on the surface of a nylon cushion layer 5, flows out from the top of the adsorption column 6 to an automatic collection sampler 9 for sampling collection, the outflow flow is regulated by a regulating valve 8, and vanadium in desorption effluent is tested every 0.5BV (V 2 O 5 Metering), desorbing until vanadium is not detected in the desorption effluent, adding 2-3 BV alkaline desorbing agent, and stopping desorbing to obtain primary vanadium-lean desorption liquid, vanadium-rich desorption liquid and secondary vanadium-lean desorption liquid in sequence. In the present invention, the flow rate of the alkaline desorbent into the sulfuric acid type ion exchange resin column (i.e., the inlet flow rate) is preferably 0.1 to 0.2mL/min, more preferably 0.15 to 0.18mL/min. In the invention, the mechanism of the desorption is shown as a formula (2) and a formula (3), wherein R represents macroporous basic anion exchange resin.
R 2 H 2 V 10 O 28 4- +2OH - ＝2ROH - +H 2 V 10 O 28 4- Formula (2).
R 2 H 2 V 10 O 28 4- +2Cl - ＝2RCl - +H 2 V 10 O 28 4- Formula (3).
In the present invention, V is 2 O 5 The concentration of the vanadium-rich desorption liquid is preferably 60-65 g/L.
After red vanadium and vanadium-rich desorption liquid are obtained, the invention uses the vanadium-rich desorption liquid to carry out alkali dissolution on the red vanadium to obtain vanadium-rich liquid.
In the invention, the ratio of the volume of the vanadium-rich desorption liquid to the dry weight of red vanadium (namely the liquid-solid ratio) is preferably 6-8 mL:1g, more preferably 7 to 8mL:1g. In the present invention, the temperature of the alkali is preferably 90 to 100 ℃, more preferably 95 ℃; the alkali dissolution time is preferably 2 to 3 hours, more preferably 2 to 2.5 hours; the pH value of the alkali is preferably 13-14, more preferably 13.5; the pH is preferably adjusted by sodium hydroxide.
After the alkali dissolution, the invention preferably further comprises the steps of cooling the obtained alkali dissolution reaction liquid to room temperature and then carrying out solid-liquid separation to obtain a liquid component and a solid component; and washing the liquid component by water, and combining the obtained washing liquid with the liquid component to obtain the vanadium-rich liquid. The cooling method is not particularly limited, and a cooling method well known to those skilled in the art, such as natural cooling, may be used. The solid-liquid separation is not particularly limited, and may be performed by a solid-liquid separation method known to those skilled in the art, such as filtration, suction filtration, or centrifugal separation. The invention has no special limit to the water consumption of the water washing, so that the medium vanadium (in V 2 O 5 Based on the total weight of the composition) is 70-90 g/L.
After the vanadium-rich liquid is obtained, the pH value of the vanadium-rich liquid is regulated to 8-9 by sulfuric acid, and then the vanadium-rich liquid is mixed with ammonium chloride, and ammonium salt vanadium precipitation is carried out, so that ammonium metavanadate is obtained.
In the present invention, the sulfuric acid is preferably dilute sulfuric acid, and the concentration of the dilute sulfuric acid is preferably 10 to 30wt%, more preferably 20wt%.
In the present invention, the ammonia addition factor of the ammonium chloride is preferably 3 to 4, more preferably 3 to 3.5.
In the invention, the temperature of the ammonium salt vanadium precipitation is preferably 90-100 ℃, more preferably 95 ℃; the time for precipitating the vanadium from the ammonium salt is preferably 1 to 2 hours, more preferably 1.5 hours.
After the ammonium salt is used for precipitating vanadium, the method preferably further comprises the step of drying the obtained wet ammonium metavanadate to obtain ammonium metavanadate. In the present invention, the drying temperature is preferably 20 to 30 ℃, more preferably room temperature; the drying time is not particularly limited, and the drying time is required to be constant.
The invention also provides a method for preparing vanadium pentoxide by utilizing the vanadium-containing metallurgical waste residues, which comprises the following steps: calcining the ammonium metavanadate prepared by the method of the technical scheme to obtain vanadium pentoxide.
In the present invention, the temperature of the calcination is preferably 500 to 600 ℃, more preferably 550 ℃, and the time of the calcination is preferably 2 to 3 hours, more preferably 2.5 hours. In the invention, the grade of the vanadium pentoxide is more than 99 percent.
The technical solutions of the present invention will be clearly and completely described in the following in connection with the embodiments of the present invention. It will be apparent that the described embodiments are only some, but not all, embodiments of the invention. All other embodiments, which can be made by those skilled in the art based on the embodiments of the invention without making any inventive effort, are intended to be within the scope of the invention.
In the following examples and comparative examples of the present invention:
the sulfuric acid type ion exchange resin is obtained by pretreating macroporous alkaline anion exchange resin at room temperature, and the pretreatment sequentially comprises: soaking in tap water for 24 hours, and washing with tap water until the resin has no peculiar smell and the washing water has no foam; soaking in ethanol for 12h, and washing with distilled water until the pH value of the washing liquid is 6-8; soaking for 12 hours by using 0.5mol/L hydrochloric acid solution, and flushing with distilled water until the pH value of the water flushing liquid is 6-8; soaking for 12 hours by using 0.5mol/L sodium hydroxide solution, and flushing with distilled water until the pH value of the water flushing liquid is 6-8; soaking in 30g/L sulfuric acid solution for 30h. The packed volume of the sulfuric acid type ion exchange resin in the sulfuric acid type ion exchange resin column (Φ8mm×250mm) was 10mL. Wherein the macroporous alkaline anion exchange resin is D231-YT macroporous strong alkaline anion exchange resin, D301 macroporous weak alkaline styrene anion exchange resin or D314 macroporous weak alkaline acrylic anion exchange resin. The 201X 7 gel-type strongly basic anion resin was also treated in accordance with the pretreatment described above to obtain a sulfuric acid-type 201X 7 gel-type strongly basic anion resin.
The main chemical components of the adopted vanadium-containing metallurgical waste residues are shown in table 1.
Adsorption-desorption (i.e., ion exchange) is performed using the column-type dynamic adsorption-desorption apparatus shown in fig. 1, and adsorption and desorption are both performed in a new column-type dynamic adsorption-desorption apparatus.
Example 1
Parameter screening for recovering vanadium in precipitation mother liquor after hydrolysis precipitation by ion exchange method
(1) Sulfuric acid curing-water leaching: crushing vanadium-containing metallurgical waste residue to a particle size of-0.125 mm, wherein the mass of the vanadium-containing metallurgical waste residue accounts for more than 90% of the total mass of the vanadium-containing metallurgical waste residue (abbreviated as-0.125 mm), adding 98wt% concentrated sulfuric acid at room temperature, curing for 3 hours under stirring, adding water, leaching for 1.5 hours under stirring at room temperature, filtering to obtain a liquid component and a solid component, washing the solid component with water, and combining the obtained washing liquid with the liquid component to obtain a vanadium-containing sulfuric acid leaching solution (pH of about 1 and V 2 O 5 The vanadium leaching rate is 99.1 percent; wherein, the concentration is H 2 SO 4 The mass of the waste slag is 70% of that of the vanadium-containing metallurgical slag, and the liquid-solid ratio of the vanadium-containing metallurgical slag to water for water immersion is 4:1.
(2) Hydrolytic precipitation of vanadium (red vanadium): adding heavy calcium carbonate into the vanadium-containing sulfuric acid leaching solution to adjust the pH value to 1.6, filtering, adding NaOH into the obtained sulfate radical-removing vanadium-containing sulfuric acid leaching solution to adjust the pH value to 2.2, and adding NaClO 3 Uniformly mixing, carrying out oxidation-vanadium precipitation reaction for 1h under the condition of stirring at 95 ℃, filtering to obtain red vanadium (vanadium precipitation rate is 80%, vanadium grade is 50%) and a liquid component, washing the solid component with water, combining the obtained washing liquid with the liquid component, and then regulating pH values (respectively 1.6, 1.8, 2.0, 2.2, 2.24 and 2.4) by using sodium hydroxide to obtain precipitation mother liquor; the mass concentration of each component in the precipitation mother liquor is shown in table 2:
TABLE 2 mass concentration of each component in precipitation mother liquor
(1) Influence of resin species on vanadium recovery
4 parts of a precipitation mother solution (pH=2.24), each 100mL of the precipitation mother solution was placed in a 250mL Erlenmeyer flask, and sulfuric acid type resins (201X 7, D231YT, D314, D3) were added respectively01 1mL each, putting into a constant temperature oscillator (25 ℃) to oscillate under 160r/min, sampling and analyzing V after adsorption balance 2 O 5 Concentration, calculating equilibrium adsorption capacity of resin, screening out resin type by comparison of adsorption exchange capacity, and comparing V with 4 sulfuric acid type ion exchange resins 2 O 5 The saturated adsorption amounts of (2) are shown in Table 3:
table 3 4 sulfuric acid ion exchange resin pairs V 2 O 5 Saturated adsorption capacity of (2)
As is clear from Table 3, in the vanadium-containing feed solution V 2 O 5 The 201X 7 gel resin has obviously weaker vanadium adsorption capacity than macroporous resin under the static adsorption condition with the concentration of 2.86g/L and the pH=2.24 and the adsorption temperature of 25 ℃. Wherein the adsorption capacity of the D314 resin is the largest, and the D314 resin is selected for separation and recovery of vanadium in the precipitation mother liquor in the subsequent test.
(2) Influence of pH on resin saturated adsorption Capacity
The existing state of vanadium in the solution is very complex, and the vanadium presents different polymerization forms along with the change of pH, and the valence state and the hydrated ion radius are also different, so that the ion exchange and the adsorption are affected. Because the concentration of iron ions in the precipitation mother liquor is high, an iron precipitation effect on the adsorption process can be generated when the pH is greater than 2.4.
Carrying out a static adsorption test according to the method of the step (1), wherein the difference from the step (1) is that the pH value of the precipitation mother liquor is respectively 1.6, 1.8, 2.0, 2.2 and 2.4, and the resin is sulfuric acid type D314 resin; by adopting a static adsorption test, the pH value is equal to that of sulfuric acid type D314 resin to V 2 O 5 The effect of the saturated adsorption amount of (2) is shown in FIG. 2.
As can be seen from fig. 2, the resin adsorption capacity increases with increasing pH in the range of 1.6 to 2.4, and the maximum resin adsorption capacity at ph=2.4 is 230mg/mL. Pentavalent vanadium in precipitation mother liquor (sulfuric acid medium) for VO 4 3- 、HV 10 O 28 5- 、H 2 V 10 O 28 4- And VO (Voice over Internet protocol) 2 SO 4 - And the main existence forms of vanadium ions in the solution at different pH values are different. When the pH value is 1.6, the pentavalent vanadium in the precipitation mother liquor is more than H 2 V 10 O 28 4- And a small amount of VO 4 3- 、VO 2 SO 4 - The ionic form exists, and at the moment, the pH value of the liquid is increased, which is favorable for H 2 V 10 O 28 4- The adsorption rate gradually increases. The recovery rate of vanadium is optimal at a pH of the precipitation mother liquor of 2.4 due to the influence of iron ion precipitation.
(3) Influence of adsorption temperature on saturated adsorption capacity of resin
In the ion exchange process, the temperature affects the swelling of the resin, the solvation of the ions, the dissociation of ion pairs or complexes in the resin, etc., thereby affecting the equilibrium concentration of adsorption of ions by the resin. Adsorption is carried out according to step (2) at 5 ℃, 15 ℃, 25 ℃, 35 ℃, 45 ℃ and 55 ℃ respectively under the condition that the pH of the precipitation mother liquor=2.4, and the temperature is equal to V 2 O 5 The effect of the saturated adsorption amount of (2) is shown in FIG. 3. As can be seen from FIG. 3, as the adsorption temperature increases, the adsorption amount of vanadium by the resin gradually increases, and the adsorption amount is 243mg/mL at the maximum at 45 ℃, and at this time, the adsorption amount does not increase again when the temperature increases. The adsorption reaction of the sulfuric acid type D314 resin on vanadium is shown to be an endothermic reaction. The temperature rise can accelerate the internal ion exchange and the external particle diffusion speed, promote the ion exchange reaction, and is also beneficial to the fracture of old bonds and the formation of new bonds, but after the adsorption reaches saturation, the temperature is too high, so that unnecessary energy consumption can be increased, and the stability of the resin can be damaged. Accordingly, the adsorption temperature is preferably in the range of 40 to 45 ℃.
(4) Influence of adsorption time on saturated adsorption capacity of resin
Ion exchange requires a certain time to reach exchange equilibrium, adsorption time has a great influence on adsorption capacity of the resin, and adsorption time (0.5, 1, 1.5, 2, 2.5, 3, 6, 9, 12, 24, 30, 48 and 52h respectively) is examined for V according to step (2) under the conditions that ph=2.4 of precipitation mother liquor and adsorption temperature is 45 ℃ 2 O 5 Is saturated withThe effect of the amount is shown in figure 4. As can be seen from FIG. 4, sulfuric acid type D314 resin vs. V with increasing adsorption time 2 O 5 The adsorption quantity of the catalyst is gradually increased, the adsorption reaches equilibrium after about 40 hours, and the saturated adsorption capacity of the resin is 243mg/mL; resin pair V 2 O 5 The adsorption of (2) is faster between 0 and 10 hours, and then gradually slows down until equilibrium. Negatively charged SO on resin reactive groups throughout the ion exchange process 4 2- H in ion-quilt solution 2 V 10 O 28 4- Ion exchange. At the very beginning of the exchange process, the resin surface SO 4 2- Ions and H in solution 2 V 10 O 28 4- The concentration difference is larger, the adsorption reaction driving force is larger, the adsorption reaction rate is higher, and the exchange sites on the resin are gradually replaced by H along with the extension of the adsorption exchange reaction time 2 V 10 O 28 4- And filling ions, wherein the mass transfer driving force is gradually reduced until the dynamic adsorption equilibrium is reached at about 40 h.
(5) Ion exchange column type adsorption
Placing the precipitation mother liquor (pH=2.24) in a feed liquid storage tank 1, pumping the precipitation mother liquor into a constant pressure tank 3 by a delivery pump 2, flowing out from the bottom opening of the constant pressure tank 3, flowing into the adsorption column 6 from the bottom at a flow rate of 0.175mL/min, flowing through sulfuric acid type ion exchange resin 7 from bottom to top, flowing out from the top of the adsorption column 6 at a flow rate of 1BV/h into an automatic collection sampler 9 for sampling and collection, regulating the outflow flow rate by a regulating valve 8, and testing the V in the adsorption effluent every 1BV 2 O 5 When the concentration of vanadium in the effluent is equal to the concentration of V in the precipitation mother liquor 2 O 5 After the concentration of the water is equal, continuing to adsorb 2-3 BV, wherein the sulfuric acid type ion exchange resin is adsorbed and saturated at the moment, then draining the column inner liquid in the sulfuric acid type ion exchange resin column, and then washing with 1BV water to obtain a saturated resin column and a column inner residual liquid respectively; and combining the residual liquid in the column with the precipitation mother liquid which is not adsorbed, and continuing the next adsorption, wherein the adsorption temperature is room temperature. Column type dynamic adsorption V 2 O 5 The curve is shown in fig. 5. As can be seen from FIG. 5, the outflow is controlledWhen the liquid speed is 0.9-1.1 BV/h, the vanadium adsorption of D314 resin is regular, the curve is S-shaped, vanadium is adsorbed at first, then suction leakage occurs, a penetration point occurs at 8-12 BV, the adsorption rate is obviously reduced until saturation, the penetration bed volume is 50BV, the saturation bed volume is 100BV, the ratio (penetration ratio) of the penetration bed volume to the saturation bed volume is 1:2, the saturated resin V capacity can reach 228mg/mL, and V 2 O 5 The adsorption rate of (2) is more than 95.5%.
(6) Ion exchange dynamic desorption
Placing 5% NaOH-10% NaCl mixed solution into a feed liquid storage tank 1, pumping into a constant pressure tank 3 by a delivery pump 2, allowing an alkaline desorbing agent to flow out from the bottom opening of the constant pressure tank 3, allowing the alkaline desorbing agent to flow into the bottom opening of an adsorption column 6 at a flow rate of 0.175mL/min, allowing the alkaline desorbing agent to flow out from the top of the adsorption column 6 into an automatic collecting sampler 9 for sampling and collecting, regulating the flow rate by a regulating valve 8, testing the concentration of vanadium in the desorbed effluent every 0.5BV, combining 2 or 3 samples after peak value into 1 sample according to the condition of the desorbed sample until the effluent does not contain vanadium, stopping desorbing after adding 2-3 BV alkaline desorbing agent, taking desorbed resin to analyze the vanadium content, allowing the desorption temperature to be room temperature, and measuring the V in the effluent by a ferrous ammonium sulfate direct titration method 2 O 5 Concentration. Column type dynamic desorption V 2 O 5 The curve is shown in fig. 6. As can be seen from FIG. 6, the desorption curve is very regular, starting with effluent V 2 O 5 The concentration is low, the concentration rises rapidly, and after reaching the peak value, the concentration gradually and gradually decreases, and the desorption is completed by using 9BV desorbent. Peak value V of desorption solution 2 O 5 Concentration 67.5g/L, vanadium enrichment ratio 23.6 times, lean resin residual V 2 O 5 The amount is about 3mg/mL, and the desorption rate is more than 98.2 percent. The desorption liquid can be divided into three sections, the first section is one-time vanadium-lean desorption liquid with the quantity of 1.5BV and V 2 O 5 The concentration is low; the second stage is vanadium-rich desorption liquid with the amount of 2.5BV and V 2 O 5 The concentration is 62.5g/L; the third stage is a secondary vanadium-lean desorption liquid with the quantity of 5BV and V 2 O 5 The concentration is low. According to V in the three-stage solution 2 O 5 Concentration analysis, the high concentration of the vanadium-rich desorption liquid can be used for dissolving red vanadium, and the vanadium concentration in the primary vanadium-lean desorption liquid and the secondary vanadium-lean desorption liquidLow, can be used for desorbing new saturated resin columns.
Example 2
The process flow chart shown in fig. 7 is adopted to prepare ammonium metavanadate and vanadium pentoxide, and the specific steps are as follows:
(1) Sulfuric acid curing-water leaching: crushing vanadium-containing metallurgical waste residue to a particle size of-0.125 mm, wherein the mass of the vanadium-containing metallurgical waste residue accounts for more than 90% of the total mass of the vanadium-containing metallurgical waste residue (abbreviated as-0.125 mm), adding concentrated sulfuric acid with a concentration of 98wt% at room temperature, curing for 3 hours under stirring, adding water, leaching for 1.5 hours under stirring at room temperature, filtering to obtain a liquid component and a solid component (sulfuric acid curing-leaching residue), washing the solid component with water, and combining the obtained water washing liquid with the liquid component to obtain vanadium-containing sulfuric acid leaching liquid (pH is about 1 and V 2 O 5 The vanadium leaching rate is 99.1 percent; wherein the mass of the concentrated sulfuric acid is 70% of the mass of the vanadium-containing metallurgical waste residue, and the liquid-solid ratio of water leaching is 4:1.
(2) Hydrolytic precipitation of vanadium (red vanadium): adding heavy calcium carbonate into the vanadium-containing sulfuric acid leaching solution to adjust the pH value to 1.6, filtering, adding NaOH into the obtained sulfate radical-removing vanadium-containing sulfuric acid leaching solution to adjust the pH value to 2.2, and adding NaClO 3 Uniformly mixing, carrying out oxidation-vanadium precipitation reaction for 1h under the condition of stirring at the temperature of 95 ℃, filtering to obtain red vanadium (the vanadium precipitation rate is 80%, the vanadium grade is 50%) and a liquid component, washing the solid component with water, combining the obtained washing liquid with the liquid component, and then regulating the pH value to 2.4 by using sodium hydroxide to obtain precipitation mother liquor; in V form 2 O 5 Calculating, the sulfate radical removal vanadium-containing sulfuric acid leaching solution and NaClO 3 The mass ratio of (2) is 1:0.05.
(3) Ion exchange adsorption: placing the precipitation mother liquor in a feed liquid storage tank 1, pumping the precipitation mother liquor into a constant pressure tank 3 by using a delivery pump 2, allowing the precipitation mother liquor to flow out from the bottom opening of the constant pressure tank 3, allowing the precipitation mother liquor to enter from the bottom of an adsorption column 6 at a flow rate of 0.175mL/min, allowing the precipitation mother liquor to flow through a sulfuric acid type ion exchange resin 7 from bottom to top, allowing the precipitation mother liquor to flow out from the top of the adsorption column 6 at a flow rate of 1BV/h into an automatic collection sampler 9 for sampling collection, regulating the flow rate by a regulating valve 8, and detecting and adsorbing vanadium (in V at intervals of 1BV 2 O 5 Meter) concentration ofVanadium (in V) 2 O 5 Calculated by V) and the concentration of vanadium (expressed as V) in the precipitation mother liquor 2 O 5 Meter), continuing to adsorb for 2-3 BV after the concentration is equal, wherein the sulfuric acid type ion exchange resin is adsorbed and saturated at the moment, draining the liquid in the sulfuric acid type ion exchange resin column, and then washing with 1BV water to obtain saturated resin column and residual liquid in the column respectively; combining the residual liquid in the column with the non-adsorbed precipitation mother liquor, and continuing the next adsorption, wherein the adsorption temperature is 45 ℃;
then placing 5% NaOH-10% NaCl mixed solution in a feed liquid storage tank 1, pumping into a constant pressure tank 3 by a delivery pump 2, flowing out alkaline desorber from the bottom opening of the constant pressure tank 3, flowing into an adsorption column 6 from the bottom at a flow rate of 0.175mL/min, flowing out from the top of the adsorption column 6 into an automatic collection sampler 9 for sampling and collection, regulating the flow rate by a regulating valve 8, and testing vanadium in the desorbed effluent every 0.5BV (in V 2 O 5 Metering) and desorbing until vanadium is not detected in the desorption effluent liquid, stopping desorbing after adding 2-3 BV alkaline desorbing agent, and sequentially obtaining primary vanadium-lean desorption liquid (1.5 BV), vanadium-rich desorption liquid (2.5 BV) and secondary vanadium-lean desorption liquid (5 BV); and adjusting the pH values of the primary vanadium-lean desorption liquid and the secondary vanadium-lean desorption liquid to 13.5 by using sodium hydroxide, and then using the solution for desorption of the next batch of saturated resin columns. Vanadium in adsorption-desorption process (in V 2 O 5 Calculated) was 95.5%.
(4) Alkali dissolution of red vanadium: adjusting pH of the vanadium-rich desorption solution to 13.8 with sodium hydroxide, mixing with red vanadium, and dissolving at 95deg.C under stirring for 2 hr to obtain vanadium-rich solution (in V) 2 O 5 Vanadium dissolution rate was 99.8%); wherein the liquid-solid ratio of the dry weight of the red vanadium to the vanadium-rich desorption liquid after the pH adjustment is 8:1 (8 mL:1 g).
(6) Ammonium salt vanadium precipitation: the pH value of the vanadium-rich liquid is adjusted to 8 by using dilute sulfuric acid with the concentration of 20 weight percent, and NH is added according to the proportion that the ammonia adding coefficient is 3 4 Mixing Cl uniformly, precipitating vanadium with ammonium salt at 90deg.C under stirring for 1 hr, cooling to room temperature, vacuum filtering, and oven drying the wet ammonium metavanadate at 25deg.C to obtain ammonium metavanadate (in V) 2 O 5 The vanadium precipitation rate was 91.8%.
(7) Calcining the ammonium metameal at 550 ℃ for 2 hours to obtain vanadium pentoxide (orange yellow, vanadium grade is more than 99.1%).
In V form 2 O 5 Total yield of vanadium=99.9% × (80% +20% ×95.5%) ×99.8% ×91.8% =90.7% (V is the amount of vanadium in the vanadium-containing metallurgical slag) 2 O 5 And (5) counting.
Example 3
Influence of sulfuric acid curing-Water leaching parameters on vanadium leaching Rate
(1) Influence of concentrated sulfuric acid consumption on vanadium leaching rate: vanadium pentoxide was prepared as in example 2, differing from example 1 only in H 2 SO 4 The mass of the waste slag is 30%, 40%, 50% and 60% of the mass of the vanadium-containing metallurgical waste slag respectively.
(2) Influence of maturation time on vanadium leaching rate: vanadium pentoxide was prepared as in example 2, differing from example 1 only in the maturation times of 1h, 3h, 4h and 5h, respectively.
(3) Influence of vanadium-containing metallurgical waste residue granularity on vanadium leaching rate: vanadium pentoxide was prepared in the same manner as in example 2, except that the vanadium-containing metallurgical slag had particle sizes of-0.6 mm, -0.425mm, -0.25mm and-0.18 mm, respectively, as in example 1.
(4) Influence of the water-leaching liquid-solid ratio on the leaching rate of vanadium: vanadium pentoxide was prepared as in example 2, differing from example 1 only in the liquid-to-solid ratios of water leaching of 2:1, 3:1, 5:1 and 6:1, respectively.
(5) Influence of water immersion time on vanadium leaching rate: vanadium pentoxide was prepared as in example 2, differing from example 1 only in the water immersion times of 0.5h, 1h, 2h and 2.5h, respectively.
Sulfuric acid aging-Water leaching parameters vs vanadium (in V 2 O 5 The results of the results are shown in Table 4 and FIG. 8, wherein (a) is the influence of the amount of concentrated sulfuric acid on the leaching of vanadium, (b) is the influence of the curing time on the leaching of vanadium, (c) is the influence of the granularity of the waste slag containing vanadium on the leaching of vanadium, (d) is the influence of the water leaching liquid-solid ratio on the leaching of vanadium, and (e) is the influence of the water leaching time on the leaching of vanadium.
TABLE 4 influence of sulfuric acid maturation-Water leaching parameters on vanadium leaching Rate
As can be seen from the table 4 and the graph 8 (a), the leaching rate of vanadium is obviously improved along with the increase of the consumption of the concentrated sulfuric acid, when the consumption of the concentrated sulfuric acid is 70% of the mass of the vanadium-containing metallurgical waste residue, vanadium is basically leached out, the leaching rate reaches 98.7%, and the leaching residue V 2 O 5 The grade is 0.13 percent. The method shows that the acid consumption of the vanadium-containing metallurgical waste residue is large, the consumption of the concentrated sulfuric acid has a key influence on the leaching of vanadium, because the vanadium-containing metallurgical waste residue contains complex components and various metal oxides, metal elements such as Pb, zn, fe, V and the like participate in the reaction in the curing process, when the consumption of the concentrated sulfuric acid is low, the vanadium-containing metallurgical waste residue cannot react with the concentrated sulfuric acid completely, and H in the solution in the water leaching process + The concentration is lower and only part of the metal enters the solution in an ionic form, so that the leaching rate of vanadium is lower. In conclusion, the method is suitable for curing by selecting the concentrated sulfuric acid with the mass of 70% of the vanadium-containing metallurgical waste residues.
As is clear from Table 4 and FIG. 8 (b), the leaching rate of vanadium gradually increased with the aging time. The leaching rate of vanadium is rapidly increased within 1h of curing time, gradually slowed down within 1-3 h, reaches 99.1% after 3h, and leached slag V 2 O 5 The grade is 0.11 percent. Thereafter, the vanadium leaching rate tends to decrease smoothly or even slowly with increasing aging time. The method shows that the curing time is short, the contact effect of the acid and the vanadium-containing metallurgical waste residue in the curing process is insufficient, and the crystal lattice of the vanadium-containing phase in the vanadium-containing metallurgical waste residue is not fully destroyed. With the extension of the curing time, the contact effect of the concentrated sulfuric acid and the vanadium-containing metallurgical waste residue is gradually and fully achieved, the breaking of chemical bonds is promoted by the heat released by curing, the crystal phase structure of the vanadium ore is thoroughly destroyed, vanadium is exposed, and the low-valence vanadium is oxidized into high-valence vanadium dissolved in acid. The aging time is prolonged again, and the leaching rate is slightly reduced due to the fact that the acid is slightly volatilized. In order to fully recycle vanadium in the vanadium-containing metallurgical waste residue, the curing time is selected to be 3 hours.
As can be seen from Table 4 and FIG. 8 (c), followingThe granularity of the vanadium-containing metallurgical waste slag is reduced, the leaching rate of vanadium is slowly improved, and when the granularity of the vanadium-containing metallurgical waste slag is minus 0.125mm, the leaching rate can reach more than 99.9 percent, and vanadium in the vanadium-containing metallurgical waste slag is basically leached out. This is because the finer the mineral particles, the larger the contact area with concentrated sulfuric acid, the more easily the lattice structure is broken, and the more easily vanadium therein is oxidized and dissolved. However, because the metallurgical waste slag has smaller granularity, the metallurgical waste slag is in full contact with concentrated sulfuric acid, and most of vanadium in the metallurgical waste slag is PbZnVO 4 The (OH) form exists and can be directly leached by acid, so that the leaching rate of the vanadium-containing metallurgical waste residue can reach more than 99 percent without fine grinding. When the granularity of the vanadium-containing metallurgical waste residue is continuously reduced, although the leaching rate is slightly improved, the grinding cost is increased due to the fact that the granularity is too fine, and a certain difficulty is caused in suction filtration, so that the vanadium-containing metallurgical waste residue is directly cured and leached under the condition that the granularity is minus 0.6 mm.
As can be seen from Table 4 and FIG. 8 (d), as the water-to-solid ratio increases, the vanadium leaching rate gradually increases, and as the liquid-to-solid ratio increases to 4:1, the vanadium leaching rate reaches 99.1%, and the leaching slag V 2 O 5 The grade is 0.09%, the liquid-solid ratio is continuously increased, and the leaching rate is not obviously increased. When the liquid-solid ratio is smaller, the concentration of ore pulp is larger, the viscosity in the leaching process is stronger, so that the mass transfer resistance of the solution and the vanadium-containing metallurgical waste residue is larger, and the vanadium leaching rate is lower. When the liquid-solid ratio is increased, the mass transfer resistance of the vanadium leaching is reduced, so that the vanadium leaching process is easy to carry out, and the leaching rate of the vanadium is increased while the leaching rate of the vanadium is improved. However, when the liquid-solid ratio is too large, the concentration of vanadium in the leaching solution is reduced, which is unfavorable for the subsequent purification and enrichment process, and the treatment capacity of the tail liquid after vanadium extraction is increased. Therefore, a liquid-to-solid ratio of 4:1 is suitable.
As is clear from table 4 and fig. 8 (e), the vanadium leaching rate gradually increased with the increase of the water leaching time. The leaching rate of vanadium is fast in the beginning, the leaching rate can reach 92.9% when the leaching time is 0.5h, the leaching rate of vanadium is slowly improved between 0.5 and 1.5h, and the leaching rate of vanadium is not increased after leaching. This is because the concentration difference of the reaction is relatively large and the reaction rate is fast when the leaching of vanadium is started, so that the leaching rate at the beginning stage is obviously increased. H as vanadium leaching proceeds + Concentration gradually increasesThe method reduces that vanadium in the vanadium-containing metallurgical waste residue is gradually leached, the mass transfer driving force in the leaching process is insufficient, and the diffusion process is slowed down, so that the leaching rate of the vanadium is gradually slowed down until the reaction is finished. When the leaching time is 1.5h, the leaching rate of vanadium can reach 99.1 percent, and the leaching slag V is leached 2 O 5 Grade 0.09%. Thus, 1.5h was chosen as the optimal leaching time.
The XRF analysis results of the sulfuric acid curing-water leaching slag obtained after sulfuric acid curing-water leaching under the conditions that the dosage of concentrated sulfuric acid is 70%, the curing time is 3h, the granularity of the vanadium-containing metallurgical slag is-0.6 mm, the water leaching liquid solid ratio is 4:1 and the water leaching time is 1.5h are shown in Table 5:
TABLE 5 chemical composition of sulfuric acid ripening-water leaching residue
It can be seen from Table 5 that the vanadium and zinc in the vanadium-containing metallurgical slag are substantially totally leached, the iron is partially leached, and the lead is enriched.
Comparative example 1
Crushing vanadium-containing metallurgical waste residues to a particle size of-0.125 mm, wherein the mass of the vanadium-containing metallurgical waste residues accounts for more than 90% of the total mass of the vanadium-containing metallurgical waste residues (abbreviated as-0.125 mm), adding a sulfuric acid solution with the concentration of 240g/L at room temperature, carrying out acid leaching for 2h under the stirring condition, carrying out suction filtration to obtain a liquid component and a solid component, and washing the solid component with water to obtain acid leaching residues, wherein the acid leaching solution solid ratio is 4:1.
An SEM image of the vanadium-containing metallurgical slag used in the step (1) of example 2, the obtained slaked slag (obtained by suction filtration after slaking, and washing the obtained solid component with water), and the sulfuric acid slaked-water leaching slag and the acid leaching slag obtained in the comparative example 1 is shown in fig. 9, wherein (a) is the vanadium-containing metallurgical slag, (b) is the acid leaching slag, (c) is the slaked slag, and (d) is the sulfuric acid slaked-water leaching slag. As can be seen from fig. 9, the vanadium-containing metallurgical slag is in the form of a block, and the valuable metals therein are packed in the block structure; the vanadium-containing metallurgical waste residue is directly subjected to acid leaching, the crystal lattice structure of the minerals is destroyed, and the original block structure is destroyed into loose fine flakes; the outside of the cured slag obtained after curing by the concentrated sulfuric acid is wrapped by the concentrated sulfuric acid to be cementedThe inside of the vanadium-containing metallurgical slag is in a convex particle shape, which indicates that sulfuric acid is permeated into the vanadium-containing metallurgical slag to damage the internal molecular structure; the morphology of the sulfuric acid curing-water leaching slag is similar to that of the acid leaching slag, and the structure is loose, but compared with the acid leaching slag, the sulfuric acid curing-water leaching slag is loose in structure, smaller in size and in fine flake and powder particle shapes. This suggests that direct acid leaching can also destroy PbZnVO 4 The molecular structure of (OH) enables vanadium in the solution to be dissolved, and the damage to the sample structure caused by the concentrated sulfuric acid curing-water leaching treatment is more thorough, and the +3 valent vanadium in the solution is more easily stripped from the original mineral structure and further oxidized into high valent vanadium to enter the solution, so that the leaching rate of the sulfuric acid curing-water leaching vanadium is higher compared with that of the direct acid leaching.
The EDS analysis chart of the vanadium-containing metallurgical slag and the prepared sulfuric acid curing-water leaching slag used in the step (1) of example 2 is shown in FIG. 10, wherein (a) is the vanadium-containing metallurgical slag and (b) is the sulfuric acid curing-water leaching slag. As can be seen from fig. 10, the characteristic peaks of vanadium and zinc in the sulfuric acid curing-water leaching slag almost completely disappear, which indicates that vanadium and zinc in the vanadium-containing metallurgical slag are basically completely leached out; the characteristic peak portion of iron is weakened. The lead is not leached substantially and is enriched in a large amount, pbSO 4 And PbFe 3 SO 4 PO 4 (OH) 6 Is left in the leaching residue. PbZnVO is removed from zinc in vanadium-containing metallurgical waste residue 4 (OH) is present in addition to Zn 2 SiO 4 And ZnS, zinc is leached into solution as the mineral lattice is destroyed, and silicon does not react with sulfuric acid, and sulfuric acid ripening-the silicon content of the water slag is also enriched.
The foregoing is merely a preferred embodiment of the present invention and it should be noted that modifications and adaptations to those skilled in the art may be made without departing from the principles of the present invention, which are intended to be comprehended within the scope of the present invention.
Claims (9)
1. The method for preparing ammonium metavanadate from vanadium-containing metallurgical waste residues is characterized by comprising the following steps of:
(1) Mixing vanadium-containing metallurgical waste residues with concentrated sulfuric acid, curing, mixing the obtained cured material with water, and leaching with water to obtain vanadium-containing sulfuric acid leaching solution; the mass ratio of the vanadium-containing metallurgical waste residue to the concentrated sulfuric acid is 1:0.6 to 0.7; the curing time is 2-5 h;
(2) Removing sulfate radical in the vanadium-containing sulfuric acid leaching solution by using an alkaline calcium source to obtain sulfate radical-removing vanadium-containing sulfuric acid leaching solution; mixing the sulfate radical-removing vanadium-containing sulfuric acid leaching solution with sodium hydroxide and sodium chlorate to perform oxidation-vanadium precipitation reaction, and carrying out solid-liquid separation to obtain red vanadium and precipitation mother liquor respectively; the pH value of the oxidation-vanadium precipitation reaction is 2-2.4;
(3) Absorbing the precipitation mother liquor to saturation by utilizing sulfuric acid type ion exchange resin, and then desorbing by adopting an alkaline desorbing agent to sequentially obtain primary vanadium-lean desorption liquid, vanadium-rich desorption liquid and secondary vanadium-lean desorption liquid; adjusting the pH values of the primary vanadium-lean desorption liquid and the secondary vanadium-lean desorption liquid to 13-14 by using sodium hydroxide, and then desorbing the saturated resin in the next batch; the volume ratio of the primary vanadium-lean desorption liquid to the secondary vanadium-lean desorption liquid is 1-1.5: 1.5 to 3.5:3.5 to 6; the alkaline desorbent is a sodium hydroxide-sodium chloride mixed solution;
(4) Alkali dissolution is carried out on the red vanadium by utilizing the vanadium-rich desorption liquid to obtain vanadium-rich liquid;
(5) And regulating the pH value of the vanadium-rich liquid to 8-9 by utilizing sulfuric acid, mixing with ammonium chloride, and carrying out ammonium salt vanadium precipitation to obtain ammonium metavanadate.
2. The method according to claim 1, wherein in step (1), the ratio of the mass of the vanadium-containing metallurgical slag to the volume of water is 1g: 3-6 mL;
the water immersion time is 1-2.5 h.
3. The method according to claim 1, wherein in the step (2), the mass ratio of the vanadium-containing sulfuric acid leaching solution to sodium chlorate is 1:0.02 to 0.1, wherein the mass of the vanadium-containing sulfuric acid leaching solution is expressed as V 2 O 5 Counting;
the temperature of the oxidation-vanadium precipitation reaction is 90-100 ℃ and the time is 1.5-2 h.
4. The method of claim 1, wherein in step (3), the sulfuric acid type ion exchange resin comprises at least one of a sulfuric acid type D231-YT macroporous strongly basic anion exchange resin, a sulfuric acid type D301 macroporous weakly basic styrenic anion exchange resin, and a sulfuric acid type D314 macroporous weakly basic acrylic anion exchange resin.
5. The method according to claim 1, wherein in the step (3), the concentration of sodium hydroxide in the sodium hydroxide-sodium chloride mixed solution is 5 to 7.5wt% and the concentration of sodium chloride is 8 to 10wt%.
6. The method according to claim 1, wherein in step (4), the ratio of the volume of the vanadium-rich desorption liquid to the dry weight of red vanadium is 6 to 8mL:1g;
the alkali dissolution temperature is 90-100 ℃, the time is 2-3 h, and the pH value is 13-14.
7. The method according to claim 1, wherein in step (5), the ammonia addition factor of the ammonium chloride is 3 to 4;
the temperature of the ammonium salt for precipitating vanadium is 90-100 ℃ and the time is 1-2 h.
8. The method for preparing vanadium pentoxide by utilizing the vanadium-containing metallurgical waste residues is characterized by comprising the following steps of:
calcining the ammonium metavanadate prepared by the method of any one of claims 1-7 to obtain vanadium pentoxide.
9. The method of claim 8, wherein the calcination is carried out at a temperature of 500 to 600 ℃ for a time of 2 to 3 hours.
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| CN101450814A (en) * | 2007-12-07 | 2009-06-10 | 湖南创大冶金集团有限公司 | Novel method for extracting vanadic anhydride from stone coal vanadium ore |
| CN102167400A (en) * | 2011-03-18 | 2011-08-31 | 中南大学 | Method for preparing vanadium pentoxide from vanadium-containing solution |
| CN102910676A (en) * | 2012-11-14 | 2013-02-06 | 西北有色地质研究院 | Preparation method of high-purity vanadium pentoxide |
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| CN105039746A (en) * | 2015-08-28 | 2015-11-11 | 长沙矿冶研究院有限责任公司 | Method for directly extracting high-purity vanadium pentoxide from stone coal vanadium ore |
| CN110994061A (en) * | 2019-10-29 | 2020-04-10 | 大连博融新材料有限公司 | Method for recovering vanadium electrolyte |
| CN114231732A (en) * | 2021-12-20 | 2022-03-25 | 攀枝花市阳润科技有限公司 | Method for deeply extracting vanadium from vanadium-containing slurry |
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| CN101450814A (en) * | 2007-12-07 | 2009-06-10 | 湖南创大冶金集团有限公司 | Novel method for extracting vanadic anhydride from stone coal vanadium ore |
| CN102167400A (en) * | 2011-03-18 | 2011-08-31 | 中南大学 | Method for preparing vanadium pentoxide from vanadium-containing solution |
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| CN110994061A (en) * | 2019-10-29 | 2020-04-10 | 大连博融新材料有限公司 | Method for recovering vanadium electrolyte |
| CN114231732A (en) * | 2021-12-20 | 2022-03-25 | 攀枝花市阳润科技有限公司 | Method for deeply extracting vanadium from vanadium-containing slurry |
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