WO2010057412A1 - A production method of vanadium oxide using extraction - Google Patents

A production method of vanadium oxide using extraction Download PDF

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Publication number
WO2010057412A1
WO2010057412A1 PCT/CN2009/074693 CN2009074693W WO2010057412A1 WO 2010057412 A1 WO2010057412 A1 WO 2010057412A1 CN 2009074693 W CN2009074693 W CN 2009074693W WO 2010057412 A1 WO2010057412 A1 WO 2010057412A1
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Prior art keywords
vanadium
extraction
vanadium oxide
production method
water
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PCT/CN2009/074693
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French (fr)
Inventor
Yi Peng
Wu BIAN
Jing Wang
Xiaojiang Wang
Ping Pan
Yiping Zhou
Wuhan Liu
Fan Zhang
Daihua Liao
Chaohui Sun
Zongquan Zhou
Mingfu Peng
Ying Wang
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Pangang Group Research Institute Co. Ltd.
Panzhihua Iron & Steel (Group) Corporation
Panzhihua New Steel & Vanadium Co., Ltd.
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Application filed by Pangang Group Research Institute Co. Ltd., Panzhihua Iron & Steel (Group) Corporation, Panzhihua New Steel & Vanadium Co., Ltd. filed Critical Pangang Group Research Institute Co. Ltd.
Priority to NZ587968A priority Critical patent/NZ587968A/en
Publication of WO2010057412A1 publication Critical patent/WO2010057412A1/en
Priority to ZA2010/06529A priority patent/ZA201006529B/en

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    • 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
    • C01—INORGANIC CHEMISTRY
    • C01P—INDEXING SCHEME RELATING TO STRUCTURAL AND PHYSICAL ASPECTS OF SOLID INORGANIC COMPOUNDS
    • C01P2006/00—Physical properties of inorganic compounds
    • C01P2006/80—Compositional purity

Definitions

  • the present invention relates to production method of vanadium oxide using extraction, and belongs to the technical field of vanadium oxide extraction.
  • Conventional vanadium extraction process using sodium salt comprises: adopting common sodium salts, such as Na2CO3, Na 2 SO 4 , or NaCl, as additives, and roasting the sodium salt with vanadium-containing raw material at high temperature, wherein vanadium in the raw material is oxidized to V 5+ by oxygen in air, and then V 5+ bonds with sodium salt to generate sodium vanadate easily dissolvable in water; leaching the roasted product with water to dissolve sodium vanadate into solution, carrying out solid-liquid separation, removing major impurities such as P, Si and so on from the solution by using CaCl 2 , adding ammonium salts such as (NH 4 ) 2 SO 4 , NH 4 Cl, (NH 4 ) 2 CO 3 , OrNH 4 NO 3 etc.
  • common sodium salts such as Na2CO3, Na 2 SO 4 , or NaCl
  • ammonium salt has to be used in an amount much higher than theoretical amount during vanadium precipitation to give high quality vanadium product, such that wastewater after vanadium precipitation has high concentration of ammonia nitrogen and sodium salt and is most difficult to be treated, e.g., NH 4 + concentration usually is as high as 2,000-8,000 mg/L or even higher, and Na concentration can be above 20g/L. Therefore, wastewater treatment is the most difficult problem to be handled in the vanadium extraction process using sodium salt.
  • One scheme is to remove heavy metals, carry out ammonia removal and sodium removal treatments, and then discharge
  • the ammonia removal technique mainly comprises air stripping method, membrane separation method, magnesium ammonium phosphate precipitation method, chemical oxidation method, zeolite adsorption method, and biological nitrogen removal method
  • the sodium removal method mainly adopts concentration crystallization method.
  • the disadvantages of this scheme comprise that cost of ammonia removal treatment and sodium removal treatment is too high to be accepted by manufacturing plants, new pollution is likely to occur during the treatment, and the recovered sodium salt is sodium sulfate containing many impurities, which will release SO 2 to pollute environment upon roasting and thus is not suitable as roasting additive.
  • the other scheme is to remove heavy metals, carry out ammonia removal and sodium removal treatments, and return condensation water for circulation.
  • the difference from the first scheme is that the water of the first scheme does not circulate, but has to meet national industrial wastewater discharge standard ( ⁇ 15mg/L for ammonia nitrogen wastewater), so it is very difficult to meet the standard, and high cost is required; although the second scheme does not discharge wastewater, a large amount of energy has to be consumed to evaporate wastewater, which also has the disadvantage of high cost, and evaporated gaseous ammonia is usually not recovered.
  • vanadium extraction process using lime or limestone as roasting additive has been studied, which targets at some inherit shortcomings of vanadium extraction process using sodium salt, including (1) strict restriction on CaO content (less than 1.5%) in vanadium slag, because vanadium recovery yield drops 4.7-9% as CaO content in the slag increases 1%; (2) high cost due to large consumption amount of sodium salt and ammonium salt; and (3) environment pollution caused by large amount of sodium salt and ammonium salt in wastewater.
  • the roasted material of vanadium extraction process using calcium salt can be leached with sulfuric aid, or with carbonate or bicarbonate. Germany Patent Application No. 2324737 and British Patent Application No.
  • Tula Vanadium Plant of former Soviet Union constructed the world first (also the only one) vanadium manufacture plant adopting calcifying roasting-sulfuric acid leaching- vanadium hydrolysis precipitation process, which can give V 2 O 5 product with purity of 88-94% (92% on average) and containing main impurities of Mn, Mg, and Ca, and then V 2 O 5 product is smelted into vanadium iron also containing many impurities; therefore the manufacturer is non-competitive on international market, and mainly supplies product to domestic market, that is the main reason that other vanadium plants do not adopt this process.
  • Russian patent application Nos To improve product competivity on market, Russian patent application Nos.
  • 2001127026/02 and 96106854/02 disclosed method for producing high quality vanadium oxide from hydrolysis precipitate containing Mn, Ca, and Mg impurities.
  • the method comprises adopting NaOH to dissolve hydrolysis precipitate, then adding ammonium salt like ammonium sulfate while heating, precipitating ammonium polyvanadate, and calcining to give high quality vanadium oxide with V 2 O 5 content higher than 98%.
  • the method reduces total amount of ammonia-containing wastewater, but still requires wastewater treatment including ammonia removal and sodium removal; therefore the technical and economic difficulties of high cost and difficult ammonia nitrogen wastewater treatment still exist, additionally, vanadium recovery rate is lowered to some extent due to long treatment process and complicated procedures.
  • the object of the present invention is to provide a vanadium oxide production method which can not only obtain high quality vanadium product but also circulate and reuse vanadium extraction wastewater.
  • the vanadium oxide production method in the present invention comprises following steps including: a. mixing vanadium -containing material with additive to give mixed material, wherein the additive is CaO or limestone, and the amount of the additive makes the CaOAV 2 Os weight ratio in the mixed material be 0.5-1.4 : 1; b. roasting the mixed material at 860 ° C-950 ° C in oxidizing atmosphere for 60-240min to give roasted material; c. adding water into the roasted material to give slurry, stirring, and slowly adding sulfuric acid solution to leach the slurry while controlling pH at 2.5-3.5; d.
  • step f removing P, Mn, and Mg impurities in the wastewater generated in step f to make Mn 2+ and Mg 2+ concentrations lower than 5g/L respectively, and P concentration lower than 0.005g/L to give circulation water, and returning the circulation water to step c for preparing slurry and to step d for washing the residue, wherein the solid raw material used in the above steps has total alkali metal amount not more than 0.3wt%, and total amount of Cl and NO 3 not more than 0.1 wt%; and the liquid raw material used in the above steps has total alkali metal amount not more than 0.1 g/L, and total amount of Cl and NO 3 not more than 0.1 g/L.
  • wastewater can be circulated and reused; and in addition, as the inventive method adopts calcifying roasting and sulfuric acid leaching to give vanadium solution substantially free of alkali metal ions, and various raw materials (including supplement water) are substantially free of easily soluble ions of alkali metal, halogen, and nitrate, balance of the easily soluble ions can be maintained during circulation process, wastewater can be circulated and reused after treatment, and thus problems of wastewater treatment of conventional vanadium extraction process using sodium salt are avoided.
  • the vanadium -containing material can be various vanadium-containing raw materials useful for preparing vanadium oxide, such as vanadium slag or other vanadium -containing raw materials (such as vanadium-containing magnetite, and vanadium lead zinc ore, etc.), preferably vanadium slag.
  • the mixed material is pulverized to particle size of 0.1mm or less, to make vanadium be easily oxidized into V to generate vanadate.
  • the oxidizing atmosphere can be pure oxygen, air, or mixture of oxygen and inert gas.
  • the inert gas is preferably nitrogen gas.
  • appropriate roasting temperature and time can be selected within the range in the present invention, for example when the roasting temperature is high (such as 950 " C), the roasting time can be shortened (such as 60min), and when the roasting temperature is low (such as 860 "C), the roasting time can be prolonged (about 240min).
  • the roasted material is cooled and pulverized to 0.18mm or less to facilitate leaching before being prepared into slurry.
  • the slurry is prepared by stirring the roasted material with water 1.5-4 times by weight.
  • the adopted water is the washing water resulted from the residue washing, and if the washing water is insufficient, circulation water is adopted for supplement.
  • the sulfuric acid solution for leaching has concentration of 10-75%, the leaching temperature is between room temperature and 58 ° C, and the leaching time is 30-90min. More preferably, pH of the slurry is regulated to 2.8-3.3 with 32-65wt% sulfuric acid solution.
  • the washing times are preferably 5-7, the amount of water used for each washing is preferably 20-35wt% of the residue on dry basis, to maintain water balance in circulation process.
  • P-removal reagent can be adopted for P-removal treatment, as long as TWP ⁇ 1000 is satisfied.
  • Ca-removal reagent can be adopted for Ca-removal treatment, as long as [Ca 2+ ] ⁇ 0.05g/L is satisfied.
  • the extraction agent may be organic substance being capable of extracting Mn 2 ⁇ Mg 2+ and Fe 3+ at pH of 2-5, and is preferably at least one of bis-(2-ethylhexyl) phosphate, mono-(2-ethylhexyl) 2-ethylhexylphosphonate, and bis-(2,4,4-trimethylpentyl) hypophosphorous acid.
  • the extraction agent may be diluted by any common diluent such as 260# solvent oil or sulfonated kerosene prior to use.
  • the extraction agent is previously saponified by ammonia to convert most exchangeable groups in the extraction agent into NH 4 + while the remaining keeps to be H + .
  • the extraction agent is previously saponified by ammonia such that the raffinate has pH of 2.8-3.8.
  • Alkaline of alkali metals such as sodium hydroxide or potassium hydroxide can not be used for saponification.
  • the organic phase containing Mn ,Mg and Fe can be subjected to reverse extraction using 2-3M sulfuric acid solution to reverse extract the Mn + ,Mg + impurity into the sulfuric acid solution such that the organic phase is regenerated and can be reused.
  • the solution resulted from the reverse extraction can be used for reverse extraction of Mn and can be used for Mn recovery Mn with wastewater neutralization residue after the reverse extraction ability decreases.
  • the wastewater can be treated according to various routine methods to remove P, Mn, and Mg, for example, lime milk is adopted to neutralize wastewater to pH 9-11, the wastewater is filtered to remove main impurities such as Mn, P, and Mg while obtaining wastewater neutralization residue useful as raw material for Mn recovery; or Mn can be individually recovered with other reagents, and then impurities like Mn, P and Mg are removed.
  • the returned circulation water may contain a certain amount OfNH 4+ , if the weight ratio of NHVMn in the P-removed and Ca-removed leachate already satisfies above requirement, extraction is not added, and vanadium precipitation is directly carried out.
  • the present invention has following beneficial effects:
  • the inventive method adopts calcifying roasting and sulfuric acid leaching technique to give vanadium solution substantially free of alkali metal ions, and various raw materials (including supplement water) are substantially free of easily soluble ions of alkali metal, halogen, and nitrate, so that balance of the easily soluble ions can be maintained during circulation, and problems of wastewater treatment of conventional vanadium extraction process using sodium salt are avoided.
  • Extraction is adopted to substitute impurities like Mn, P and Mg with NH 4 + to refine the vanadium-containing solution so as to prepare high quality vanadium oxide product, so that technical problem that conventional calcifying roasting-sulfuric acid leaching process can not produce high quality vanadium product is solved.
  • Weight ratio of NH 3 /Mn is controlled to control the extent of the extraction, when the ratio in the leachate is lower than the value specified in the present invention, extraction is performed to regulate the ratio to the specified value, and then the vanadium precipitation is carried out; when the ratio in the leachate has already reached the specified value in the present invention, extraction is not needed, and vanadium precipitation is directly carried out, and part of NH 4 is discharged along with ammonium polyvanadate precipitate, so that NH 4 concentration in the system will not continuously increase to enable circulation and reuse.
  • the circulation water containing sulfate salt such as (NH 4 ) 2 SO 4 is adopted for leaching; within the leaching condition range in the present invention, NH 4 + will not cause adverse influence, and SO 4 " is beneficial for increasing leaching rate of calcified roasted material, so that water circulation and reuse can be finally realized. 5.
  • the inventive method can greatly increase total recovery rate (up to 82-85%) of vanadium oxide from vanadium slag, while the recovery rate of the conventional vanadium extraction process using sodium salt is about 80%; therefore the inventive method has increased the recovery rate by 2-5% on average compared with conventional vanadium extraction process using sodium salt, and the obtained vanadium product has good quality, and meets Chinese National Standard No. GB3283-87.
  • the inexpensive lime material is adopted to replace expensive sodium carbonate, the consumption amount of sulfuric acid is similar to that in the vanadium extraction process using sodium salt, and consumption and cost of other auxiliary materials are low; therefore, consumption and cost of various auxiliary and raw materials are significantly decreased.
  • Fig. 1 shows flow chart of one preferred embodiment of the method in the present invention.
  • the mixed material is calcified and roasted (equivalent to step b);
  • Sulfuric acid solution is adopted to leach the roasted material at constant pH (equivalent to step c);
  • the leachate is subjected to extraction (equivalent to step e);
  • Vanadium precipitation is performed, and ammonium polyvanadate obtained from vanadium precipitation is calcined or reduced to give high quality vanadium oxide (equivalent to step f);
  • the wastewater obtained from vanadium precipitation is added with lime milk for neutralization, the obtained circulation water is returned to prepare slurry or wash residue, and
  • Mn is recovered from the neutralization residue (equivalent to step g).
  • Example 1 Roasted material prepared according to step (1) is adopted, 51 cycles of wastewater circulation test are carried out according to the inventive method, and each cycle includes steps (2)-(6).
  • Common vanadium slag with composition shown in Table 1 is ground to less than 0.098mm, 28kg of ground vanadium slag powder is mixed with 1.96kg of lime (ground to below 0.1mm) containing CaO>98%, and then the mixture is roasted at 860 ° C in air for 240min, cooled, and ground to less than 0.18mm.
  • 2,000ml of water resulted from residue washing of last cycle (clear water is used for the first cycle) is added to 50Og of the ground roasted material to prepare slurry, 10-32wt% sulfuric acid solution is slowly added continuously while stirring, pH is controlled at 2.8-3.3 during leaching process while the slurry temperature is held at a temperature between room temperature and 58 ° C, the reaction is carried out for 60min.
  • the resultant is filtered to give leachate, the residue is washed with circulation water (clear water is used for the first cycle) for 6-7 times, water used for each time is 120ml, the washing water is combined for preparing slurry for leaching of next cycle, and the residue is dried and weighed to determine TV (total vanadium) content and calculate vanadium leaching rate.
  • the leachate is subjected to P-removal treatment and Ca-removal treatment to make P satisfy TV/P ⁇ IOOO and [Ca + ] ⁇ 0.05g/L, and then extraction is carried out.
  • Extraction 10-20volume% kerosene solution of P204 is saponified by ammonia.
  • the saponified organic phase is mixed with leachate to perform one-stage extraction at room temperature for 5min under stirring.
  • the mixture is kept standing for the phase separation, and the organic substances remained in the aqueous phase are removed
  • the weight ratio of NH 3 /Mn in the aqueous phase solution is controlled to be 0.6-50 : 1 by controlling saponification ratio and the extraction phase ratio .
  • P204 represents bis-(2-ethylhexyl) phosphate.
  • a small amount of sulfuric acid is used to regulate pH of the oil-removed aqueous phase to 1.5-2.5, then the resultant is heated to above 90 ° C, held for 60-120min, and filtered, the precipitate is washed with tap water containing [Na+K] ⁇ 0. lg/L for 3 times and water used for each time is 30ml, and the washing wastewater is combined with supernatant of vanadium precipitation to give wastewater of vanadium precipitation.
  • Ammonium polyvanadate precipitate is dried, calcined and melted at 500 ° C-800 ° C to give V 2 O 5 , and the composition of the obtained V 2 O 5 is analyzed.
  • Lime milk with low water content is prepared, added into the wastewater of vanadium precipitation to regulate pH of the solution to 9.0-10.0, and filtered.
  • the filtrate is regulated to pH of 5-7 with diluted sulfuric acid to give circulation water as residue washing water for leaching in the next cycle.
  • the neutralization residue obtained from the filtration is used as raw material for Mn recovery.
  • the organic phase containing cation such as Mn 2+ (carried organic phase) is reused after being reverse extracted by 2.5M sulfuric acid solution. After 10 cycles, the organic phase is reverse extracted by 6N hydrochloric acid besides the 2.5M sulfuric acid solution, and washed by dilute sulfuric acid to remove the Cl " , and then the extraction agent returns to the system. The solution resulted from the hydrochloric acid reverse extraction and the solution resulted from the sulfuric acid solution reverse extraction are separately treated. When Mn in the reverse-extraction solution accumulates to an extent, the reverse-extraction solution reacts with wastewater neutralization residue to recover Mn.
  • steps (2)-(6) are repeated, 50Og of roasted material is used for each cycle, liquid solid ratio for each cycle is 4:1, vanadium precipitation is not carried out after leaching in the first cycle, the leachate is used for slurry preparation of the second cycle to increase vanadium concentration of the leachate; then during leaching in each of the rest cycles, residue washing water of last cycle is used in step (2), and the insufficient part is supplemented by circulation water; circulation water is used for washing the residue, and the insufficient part is supplemented by clear water. 51 cycles are carried out, and no wastewater containing ammonia nitrogen is discharged from the system.
  • the maximum value, the minimum value, and the average value of the recovery rate of the rest cycles are shown in Table 2; the maximum value, the minimum value, and the average value of the chemical compositions of the V 2 O 5 product of the rest cycles are shown in Table 3, and the Table 3 also shows compositions of the metallurgical No. 98 and No. 99 of Chinese National Standard No. GB3283-87; and the maximum value, the minimum value, and the average value of the composition of the circulation water after circulation treatment for the rest cycles are shown in Table 4.
  • the procedure of raw material pretreatment includes pulverization of bulk coarse vanadium slag, grinding, and iron removal, and a small amount of vanadium is lost in this process.
  • the table 3 shows that the vanadium product obtained by the process has good product quality, and meets Chinese National Standard No. GB3283-87.
  • Example 2 Roasted material prepared according to step (1) is adopted, 51 cycles of wastewater circulation test are carried out according to the inventive method, and each cycle includes steps (2)-(6).
  • the vanadium slag shown in Table 5 is ground to less than 0.098mm.
  • the resultant is filtered to give leachate, the residue is washed with circulation water (clear water is used for the first cycle) for 6 times, water used for each time is 120ml, the washing water is combined for preparing slurry for leaching of next cycle, and the residue is dried and weighed to determine TV (total vanadium) content and calculate vanadium leaching rate.
  • the leachate is subjected to P-removal treatment and Ca-removal treatment to make P satisfy TV/P ⁇ IOOO and [Ca + ] ⁇ 0.05g/L, and then extraction is carried out.
  • 10-30volume% kerosene solution of P507 is saponified by ammonia.
  • the saponified organic phase is mixed with leachate to perform one-stage extraction at room temperature for 5min under stirring.
  • the mixture is kept standing for the phase separation, and the organic substances remained in the aqueous phase are removed
  • the weight ratio of NH 3 /Mn in the aqueous phase solution is controlled to be 10-200 : 1 by controlling saponification ratio and the extraction phase ratio.
  • P507 represents mono-(2-ethylhexyl) 2-ethylhexylphosphonate. (4) Vanadium precipitation and V 2 O 5 production by calcination
  • a small amount of sulfuric acid is used to regulate pH of the aqueous phase solution to 1.5-2.5, then the resultant is heated to above 90 ° C, held for 60-120min, and filtered, the precipitate is washed with tap water containing [Na+K] ⁇ 0.1g/L for 3 times and water used for each time is 30ml, and the washing wastewater is combined with supernatant of vanadium precipitation to give wastewater of vanadium precipitation.
  • Ammonium polyvanadate precipitate is dried, calcined and melted at 500 ° C-800 ° C to give V 2 O 5 , and the composition of the obtained V 2 O 5 is analyzed. (5) Wastewater treatment
  • Lime milk with low water content is prepared, added into the wastewater of vanadium precipitation to regulate pH of the solution to 9.5-10.00, and filtered.
  • the filtrate is regulated to pH of 5-7 with diluted sulfuric acid to give circulation water as residue washing water for leaching in the next cycle.
  • the neutralization residue obtained from the filtration is used as raw material for Mn recovery. (6) Regeneration of the organic phase
  • the organic phase containing cation such as Mn 2+ is reused after being reverse extracted by 2.5M sulfuric acid solution. After 10 cycles, the organic phase is reverse extracted by 6N hydrochloric acid besides the 2.5M sulfuric acid solution, and washed by dilute sulfuric acid to remove the Cl " , and then the extraction agent returns to the system.
  • the solution resulted from the hydrochloric acid reverse extraction and the solution resulted from the sulfuric acid solution reverse extraction are separately treated.
  • Mn in the reverse-extraction solution accumulates to an extent, the reverse-extraction solution reacts with wastewater neutralization residue to recover Mn.
  • steps (2)-(6) are repeated, 50Og of roasted material is used for each cycle, liquid solid ratio for each cycle is 2.5:1, vanadium precipitation is not carried out after leaching in the first cycle, the leachate is used for slurry preparation of the second cycle to increase vanadium concentration of the leachate; then during leaching in each of the rest cycles, residue washing water of last cycle is used in step (2), and the insufficient part is supplemented by circulation water; circulation water is used for washing the residue, and the insufficient part is supplemented by clear water. 51 cycles are carried out, and no wastewater containing ammonia nitrogen is discharged from the system.
  • the table 7 shows that the vanadium product obtained by the process has good product quality, and meets Chinese National Standard No. GB3283-87.
  • roasted material prepared according to step (1) is adopted, 100 cycles of wastewater circulation test are carried out according to the inventive method, and each cycle includes steps (2)-(6).
  • Preparation of roasted material 110kg of vanadium slag with composition shown in Table 9 is ground to less than 0.098mm, the ground vanadium slag powder is mixed with 7.7kg of lime (ground to below 0.1mm) containing 98% CaO, and then the mixture is roasted at 920 ° C in air for 150min, cooled, and ground to less than 0.18mm.
  • 2,000ml of water resulted from residue washing of last cycle (clear water is used for the first cycle) is added to 1,00Og of ground roasted material to prepare slurry, 50-75wt% sulfuric acid solution is slowly added continuously while stirring, pH is controlled at 2.8-3.3 during leaching process while the slurry temperature is held at a temperature between room temperature and 58 ° C, the reaction is carried out for 60min.
  • the resultant is filtered to give leachate, the residue is washed with circulation water (clear water is used for the first cycle) for 5-6 times, water used for each time is 250ml, the washing water is combined for preparing slurry for leaching of next cycle, and the residue is dried and weighed to determine TV (total vanadium) content and calculate vanadium leaching rate.
  • the leachate is subjected to P-removal treatment and Ca-removal treatment to make P satisfy TV/P ⁇ IOOO and [Ca + ] ⁇ 0.05g/L, and then extraction is carried out.
  • Cyanex represents bis-(2,4,4-trimethylpentyl) hypophosphorous acid.
  • a small amount of sulfuric acid is used to regulate pH of the aqueous phase solution to 1.5-2.5, then the resultant is heated to above 90 ° C, held for 60-120min, and filtered, the precipitate is washed with tap water containing [Na+K] ⁇ 0.1g/L for 3 times and water used for each time is 30ml, and the washing wastewater is combined with supernatant of vanadium precipitation to give wastewater of vanadium precipitation.
  • Ammonium polyvanadate precipitate is dried, calcined and melted at 500 ° C-800 ° C to give V2O5, and the composition of the obtained V2O5 is analyzed.
  • Lime milk with low water content is prepared, added into the wastewater of vanadium precipitation to regulate pH of the solution to 9.5-11, and filtered.
  • the filtrate is regulated to pH of 5-7 with diluted sulfuric acid to give circulation water as residue washing water for leaching in the next cycle.
  • the neutralization residue obtained from the filtration is used as raw material for Mn recovery.
  • the organic phase containing cation such as Mn 2+ is reused after being reverse extracted by 2.5M sulfuric acid solution. After 10 cycles, the organic phase is reverse extracted by 6N hydrochloric acid besides the 2.5M sulfuric acid solution, and washed by dilute sulfuric acid to remove the Cl " , and then the extraction agent returns to the system.
  • the solution resulted from the hydrochloric acid reverse extraction and the solution resulted from the sulfuric acid solution reverse extraction are separately treated.
  • Mn in the reverse-extraction solution accumulates to an extent, the reverse-extraction solution reacts with wastewater neutralization residue to recover Mn.
  • steps (2)-(6) are repeated, lOOOg of roasted material is used for each cycle, liquid solid ratio for each cycle is 2:1, vanadium precipitation is not carried out after leaching in the first cycle, the leachate is used for slurry preparation of the second cycle to increase vanadium concentration of the leachate; then during leaching in each of the rest cycles, residue washing water of last cycle is used in step (2), and the insufficient part is supplemented by circulation water; circulation water is used for washing the residue, and the insufficient part is supplemented by clear water. 51 cycles are carried out, and no wastewater containing ammonia nitrogen is discharged from the system.
  • the table 11 shows that the vanadium product obtained by the process has good product quality, and meets Chinese National Standard No. GB3283-87.

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  • Organic Chemistry (AREA)
  • Inorganic Chemistry (AREA)
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Abstract

The present invention relates to a production method of vanadium oxide using extraction, and belongs to the technical field of vanadium oxide extraction. The technical problem to be solved in the present invention is to provide a clean vanadium oxide production method which can not only obtain high quality vanadium product but also circulate and reuse vanadium extraction wastewater. The inventive method comprises preparing raw material to be roasted, calcifying roasting, leaching, solid-liquid separating, extraction, vanadium precipitating, and removing ammonia by calcination or reducing to prepare vanadium oxide; where vanadium extraction wastewater is returned to the system for circulation and reuse after neutralization treatment with lime milk, and no wastewater discharge is realized. The invention improves vanadium recovery rate to make it higher than that of available processes, and reduces production cost. By combining with other techniques, the invention also can convert waste resulted from extraction into secondary resource for reuse, so as to realize clean production.

Description

A production method of vanadium oxide using extraction
Technical field
The present invention relates to production method of vanadium oxide using extraction, and belongs to the technical field of vanadium oxide extraction.
Technical background
Conventional vanadium extraction process using sodium salt comprises: adopting common sodium salts, such as Na2CO3, Na2SO4, or NaCl, as additives, and roasting the sodium salt with vanadium-containing raw material at high temperature, wherein vanadium in the raw material is oxidized to V5+ by oxygen in air, and then V5+ bonds with sodium salt to generate sodium vanadate easily dissolvable in water; leaching the roasted product with water to dissolve sodium vanadate into solution, carrying out solid-liquid separation, removing major impurities such as P, Si and so on from the solution by using CaCl2, adding ammonium salts such as (NH4)2SO4, NH4Cl, (NH4)2CO3, OrNH4NO3 etc. thereto, regulating pH of the solution to 1.5-2.5, heating the solution to above 90°C , and holding at the temperature for 40-90min to separate out ammonium polyvanadate precipitate from the solution; removing impurities such as sodium salts carried in the ammonium polyvanadate precipitate by washing with water after filtering, drying, and calcining the precipitate to remove ammonia to give V2O5 , or reducing the precipitate at high temperature with reducing gas such as coal gas or natural gas to give V2O3. Presently, most manufacturers in the world adopt this process to produce vanadium products.
This process has the advantages of high product quality, stable process, and easy control. However, ammonium salt has to be used in an amount much higher than theoretical amount during vanadium precipitation to give high quality vanadium product, such that wastewater after vanadium precipitation has high concentration of ammonia nitrogen and sodium salt and is most difficult to be treated, e.g., NH4 + concentration usually is as high as 2,000-8,000 mg/L or even higher, and Na concentration can be above 20g/L. Therefore, wastewater treatment is the most difficult problem to be handled in the vanadium extraction process using sodium salt. If the wastewater is directly returned to leaching step for circulation, as sodium vanadate in roasted material (also referred to as "roasted product") is continuously dissolved in water while sodium salt can not be discharged from the solution, sodium salt concentration in the solution becomes increasingly high, the ammonium salt amount required by vanadium precipitation becomes increasingly large, the solution quickly becomes viscous, it is difficult to carry out filtration or vanadium precipitation, and actually the circulation can not be continued after 1 -2 cycles; therefore the wastewater can not be directly circulated and reused. Presently, there are mainly two schemes for solving pollution problems of wastewater from vanadium extraction process using sodium salt. One scheme is to remove heavy metals, carry out ammonia removal and sodium removal treatments, and then discharge wherein the ammonia removal technique mainly comprises air stripping method, membrane separation method, magnesium ammonium phosphate precipitation method, chemical oxidation method, zeolite adsorption method, and biological nitrogen removal method, and the sodium removal method mainly adopts concentration crystallization method. The disadvantages of this scheme comprise that cost of ammonia removal treatment and sodium removal treatment is too high to be accepted by manufacturing plants, new pollution is likely to occur during the treatment, and the recovered sodium salt is sodium sulfate containing many impurities, which will release SO2 to pollute environment upon roasting and thus is not suitable as roasting additive. The other scheme is to remove heavy metals, carry out ammonia removal and sodium removal treatments, and return condensation water for circulation. The difference from the first scheme is that the water of the first scheme does not circulate, but has to meet national industrial wastewater discharge standard ( ≤Ξ 15mg/L for ammonia nitrogen wastewater), so it is very difficult to meet the standard, and high cost is required; although the second scheme does not discharge wastewater, a large amount of energy has to be consumed to evaporate wastewater, which also has the disadvantage of high cost, and evaporated gaseous ammonia is usually not recovered.
Therefore, treatment and utilization of wastewater having high ammonia nitrogen and high sodium salt resulted from vanadium extraction process using sodium salt is urgently to be solved in the field.
Since 1960s and 1970s, vanadium extraction process using lime or limestone as roasting additive, usually called vanadium extraction process using calcium salt or calcifying roasting vanadium extraction process, has been studied, which targets at some inherit shortcomings of vanadium extraction process using sodium salt, including (1) strict restriction on CaO content (less than 1.5%) in vanadium slag, because vanadium recovery yield drops 4.7-9% as CaO content in the slag increases 1%; (2) high cost due to large consumption amount of sodium salt and ammonium salt; and (3) environment pollution caused by large amount of sodium salt and ammonium salt in wastewater. The roasted material of vanadium extraction process using calcium salt can be leached with sulfuric aid, or with carbonate or bicarbonate. Germany Patent Application No. 2324737 and British Patent Application No. 1394024 reported a method for leaching calcified roasted material with sodium carbonate solution. US Patent No. 3853985 reported a method for leaching calcified roasted material with ammonium carbonate or ammonium bicarbonate. "Thermodynamics and kinetics of vanadium slag calcium salt roasting- carbonate leaching" (Vanadium titanium, 1997 ', No. 6: 1-6) reported thermodynamics and kinetics of leaching the calcified roasted material using carbonate and bicarbonate. Due to use of sodium salt and ammonium slat, the above methods in the literatures also have the problem of ammonia nitrogen wastewater treatment. British Patent Application No. 1394024 also reported a method including leaching roasted material with sulfuric acid or hydrochloric acid, regulating pH of the acidic leachate to 1.6-1.9, heating to hydrolyze and precipitate vanadium, and drying and calcining the precipitate to give vanadium oxide product containing about 93.5% Of V2O5 and considerable amount of impurities. "Study of V2O5 extraction process by vanadium slag lime roasting method" (Iron Steel Vanadium Titanium, 1992, 13(6): 1-9) reported the study on V2O5 production by subjecting atomized vanadium slag to calcifying roasting and sulfuric acid leaching, in which the obtained acidic leachate is added with sulfuric acid to regulate pH to 2, and then heated to hydrolyze and precipitate vanadium to give product with purity of 93.6-93.92%. "Study of V2O5 extraction by vanadium slag calcium salt roasting-sulfuric acid leaching" reported study of calcifying roasting and sulfuric acid leaching of vanadium slag. Tula Vanadium Plant of former Soviet Union constructed the world first (also the only one) vanadium manufacture plant adopting calcifying roasting-sulfuric acid leaching- vanadium hydrolysis precipitation process, which can give V2O5 product with purity of 88-94% (92% on average) and containing main impurities of Mn, Mg, and Ca, and then V2O5 product is smelted into vanadium iron also containing many impurities; therefore the manufacturer is non-competitive on international market, and mainly supplies product to domestic market, that is the main reason that other vanadium plants do not adopt this process. To improve product competivity on market, Russian patent application Nos. 2001127026/02 and 96106854/02 disclosed method for producing high quality vanadium oxide from hydrolysis precipitate containing Mn, Ca, and Mg impurities. The method comprises adopting NaOH to dissolve hydrolysis precipitate, then adding ammonium salt like ammonium sulfate while heating, precipitating ammonium polyvanadate, and calcining to give high quality vanadium oxide with V2O5 content higher than 98%. The method reduces total amount of ammonia-containing wastewater, but still requires wastewater treatment including ammonia removal and sodium removal; therefore the technical and economic difficulties of high cost and difficult ammonia nitrogen wastewater treatment still exist, additionally, vanadium recovery rate is lowered to some extent due to long treatment process and complicated procedures.
Until now, there's no relevant report regarding process which can not only give high quality vanadium product, but also completely solve the problem of vanadium extraction wastewater treatment and reuse. Summary of the invention
The object of the present invention is to provide a vanadium oxide production method which can not only obtain high quality vanadium product but also circulate and reuse vanadium extraction wastewater.
The vanadium oxide production method in the present invention comprises following steps including: a. mixing vanadium -containing material with additive to give mixed material, wherein the additive is CaO or limestone, and the amount of the additive makes the CaOAV2Os weight ratio in the mixed material be 0.5-1.4 : 1; b. roasting the mixed material at 860°C-950°C in oxidizing atmosphere for 60-240min to give roasted material; c. adding water into the roasted material to give slurry, stirring, and slowly adding sulfuric acid solution to leach the slurry while controlling pH at 2.5-3.5; d. removing residue after leaching to give leachate, subjecting the leachate to P -removal treatment and Ca-removal treatment to make weight ratio of total V to P in the leachate >1000 and [Ca +] in the leachate <0.05g/L, washing the residue with circulation water to give washing water useful for preparing slurry in next leaching; e. subjecting the P -removed and Ca-removed leachate to extraction by using an extraction agent which has been saponified by ammonia, and then separating the aqueous phase and organic phase, the weight ratio of NH3AMn in the aqueous phase being controlled within 0.6-2000 : 1, the extraction agent being capable of extracting Mn2+ ,Mg2+ and Fe3+ at pH of2-5; f. regulating pH of the aqueous phase with sulfuric acid to 1.5-2.5, heating to a temperature between 90 °C and boiling temperature, holding at the temperature for 30-120min, filtering, and washing and drying the precipitate to give ammonium polyvanadate, subjecting the ammonium polyvanadate to calcination to remove ammonia to give V2Os, or reducing the ammonium polyvanadate to give V2O3; and g. removing P, Mn, and Mg impurities in the wastewater generated in step f to make Mn2+ and Mg2+concentrations lower than 5g/L respectively, and P concentration lower than 0.005g/L to give circulation water, and returning the circulation water to step c for preparing slurry and to step d for washing the residue, wherein the solid raw material used in the above steps has total alkali metal amount not more than 0.3wt%, and total amount of Cl and NO3 not more than 0.1 wt%; and the liquid raw material used in the above steps has total alkali metal amount not more than 0.1 g/L, and total amount of Cl and NO3 not more than 0.1 g/L. According to the present invention, wastewater can be circulated and reused; and in addition, as the inventive method adopts calcifying roasting and sulfuric acid leaching to give vanadium solution substantially free of alkali metal ions, and various raw materials (including supplement water) are substantially free of easily soluble ions of alkali metal, halogen, and nitrate, balance of the easily soluble ions can be maintained during circulation process, wastewater can be circulated and reused after treatment, and thus problems of wastewater treatment of conventional vanadium extraction process using sodium salt are avoided.
In step a, the vanadium -containing material can be various vanadium-containing raw materials useful for preparing vanadium oxide, such as vanadium slag or other vanadium -containing raw materials (such as vanadium-containing magnetite, and vanadium lead zinc ore, etc.), preferably vanadium slag. Preferably, in step a, the mixed material is pulverized to particle size of 0.1mm or less, to make vanadium be easily oxidized into V to generate vanadate.
In step b, the oxidizing atmosphere can be pure oxygen, air, or mixture of oxygen and inert gas. The inert gas is preferably nitrogen gas. During roasting, appropriate roasting temperature and time can be selected within the range in the present invention, for example when the roasting temperature is high (such as 950 "C), the roasting time can be shortened (such as 60min), and when the roasting temperature is low (such as 860 "C), the roasting time can be prolonged (about 240min).
Preferably, in step c, the roasted material is cooled and pulverized to 0.18mm or less to facilitate leaching before being prepared into slurry. Preferably, in step c, the slurry is prepared by stirring the roasted material with water 1.5-4 times by weight. The adopted water is the washing water resulted from the residue washing, and if the washing water is insufficient, circulation water is adopted for supplement. The sulfuric acid solution for leaching has concentration of 10-75%, the leaching temperature is between room temperature and 58 °C, and the leaching time is 30-90min. More preferably, pH of the slurry is regulated to 2.8-3.3 with 32-65wt% sulfuric acid solution. During washing the residue, the washing times are preferably 5-7, the amount of water used for each washing is preferably 20-35wt% of the residue on dry basis, to maintain water balance in circulation process.
P-removal reagent can be adopted for P-removal treatment, as long as TWP ^ 1000 is satisfied. Ca-removal reagent can be adopted for Ca-removal treatment, as long as [Ca2+]≤0.05g/L is satisfied.
In step e, the extraction agent may be organic substance being capable of extracting Mn2^Mg2+ and Fe3+ at pH of 2-5, and is preferably at least one of bis-(2-ethylhexyl) phosphate, mono-(2-ethylhexyl) 2-ethylhexylphosphonate, and bis-(2,4,4-trimethylpentyl) hypophosphorous acid. The extraction agent may be diluted by any common diluent such as 260# solvent oil or sulfonated kerosene prior to use.
The extraction agent is previously saponified by ammonia to convert most exchangeable groups in the extraction agent into NH4 + while the remaining keeps to be H+. Preferably, the extraction agent is previously saponified by ammonia such that the raffinate has pH of 2.8-3.8. Alkaline of alkali metals such as sodium hydroxide or potassium hydroxide can not be used for saponification.
Furthermore, in step e, after extraction, the organic phase containing Mn ,Mg and Fe can be subjected to reverse extraction using 2-3M sulfuric acid solution to reverse extract the Mn +,Mg + impurity into the sulfuric acid solution such that the organic phase is regenerated and can be reused. The solution resulted from the reverse extraction can be used for reverse extraction of Mn and can be used for Mn recovery Mn with wastewater neutralization residue after the reverse extraction ability decreases. When Fe in the leachate accumulates to an extent in the organic phase, Fe is reverse extracted by 6N hydrochloric acid, and the organic phase returns to the system after washing Cl" away. In step g, the wastewater can be treated according to various routine methods to remove P, Mn, and Mg, for example, lime milk is adopted to neutralize wastewater to pH 9-11, the wastewater is filtered to remove main impurities such as Mn, P, and Mg while obtaining wastewater neutralization residue useful as raw material for Mn recovery; or Mn can be individually recovered with other reagents, and then impurities like Mn, P and Mg are removed. As the returned circulation water may contain a certain amount OfNH4+, if the weight ratio of NHVMn in the P-removed and Ca-removed leachate already satisfies above requirement, extraction is not added, and vanadium precipitation is directly carried out. The present invention has following beneficial effects:
1. The inventive method adopts calcifying roasting and sulfuric acid leaching technique to give vanadium solution substantially free of alkali metal ions, and various raw materials (including supplement water) are substantially free of easily soluble ions of alkali metal, halogen, and nitrate, so that balance of the easily soluble ions can be maintained during circulation, and problems of wastewater treatment of conventional vanadium extraction process using sodium salt are avoided. 2. Extraction is adopted to substitute impurities like Mn, P and Mg with NH4 + to refine the vanadium-containing solution so as to prepare high quality vanadium oxide product, so that technical problem that conventional calcifying roasting-sulfuric acid leaching process can not produce high quality vanadium product is solved.
3. Weight ratio of NH3/Mn is controlled to control the extent of the extraction, when the ratio in the leachate is lower than the value specified in the present invention, extraction is performed to regulate the ratio to the specified value, and then the vanadium precipitation is carried out; when the ratio in the leachate has already reached the specified value in the present invention, extraction is not needed, and vanadium precipitation is directly carried out, and part of NH4 is discharged along with ammonium polyvanadate precipitate, so that NH4 concentration in the system will not continuously increase to enable circulation and reuse.
4. The circulation water containing sulfate salt such as (NH4)2SO4 is adopted for leaching; within the leaching condition range in the present invention, NH4 + will not cause adverse influence, and SO4 " is beneficial for increasing leaching rate of calcified roasted material, so that water circulation and reuse can be finally realized. 5. The inventive method can greatly increase total recovery rate (up to 82-85%) of vanadium oxide from vanadium slag, while the recovery rate of the conventional vanadium extraction process using sodium salt is about 80%; therefore the inventive method has increased the recovery rate by 2-5% on average compared with conventional vanadium extraction process using sodium salt, and the obtained vanadium product has good quality, and meets Chinese National Standard No. GB3283-87.
6. The inexpensive lime material is adopted to replace expensive sodium carbonate, the consumption amount of sulfuric acid is similar to that in the vanadium extraction process using sodium salt, and consumption and cost of other auxiliary materials are low; therefore, consumption and cost of various auxiliary and raw materials are significantly decreased.
Brief Description of Drawings
Fig. 1 shows flow chart of one preferred embodiment of the method in the present invention.
Preferred Embodiments
The present invention will be further described through following examples.
The following examples all adopt the flow as shown in the Fig. 1. The preferred examples are briefly described in combination with Fig. 1. Firstly, vanadium slag and lime is mixed to give mixed material (equivalent to step a);
The mixed material is calcified and roasted (equivalent to step b);
Sulfuric acid solution is adopted to leach the roasted material at constant pH (equivalent to step c);
After leaching, residue and leachate are separated out, the residue is washed with diluted sulfuric acid solution of pH 3-6, and the washing water is collected for preparing slurry
(equivalent to step d);
The leachate is subjected to extraction (equivalent to step e);
Vanadium precipitation is performed, and ammonium polyvanadate obtained from vanadium precipitation is calcined or reduced to give high quality vanadium oxide (equivalent to step f); The wastewater obtained from vanadium precipitation is added with lime milk for neutralization, the obtained circulation water is returned to prepare slurry or wash residue, and
Mn is recovered from the neutralization residue (equivalent to step g).
Example 1 Roasted material prepared according to step (1) is adopted, 51 cycles of wastewater circulation test are carried out according to the inventive method, and each cycle includes steps (2)-(6). (1) Preparation of roasted material:
Common vanadium slag with composition shown in Table 1 is ground to less than 0.098mm, 28kg of ground vanadium slag powder is mixed with 1.96kg of lime (ground to below 0.1mm) containing CaO>98%, and then the mixture is roasted at 860 °C in air for 240min, cooled, and ground to less than 0.18mm.
Table 1 Main components of vanadium slag (%)
Figure imgf000011_0001
(2) Leaching of roasted material
2,000ml of water resulted from residue washing of last cycle (clear water is used for the first cycle) is added to 50Og of the ground roasted material to prepare slurry, 10-32wt% sulfuric acid solution is slowly added continuously while stirring, pH is controlled at 2.8-3.3 during leaching process while the slurry temperature is held at a temperature between room temperature and 58°C, the reaction is carried out for 60min. The resultant is filtered to give leachate, the residue is washed with circulation water (clear water is used for the first cycle) for 6-7 times, water used for each time is 120ml, the washing water is combined for preparing slurry for leaching of next cycle, and the residue is dried and weighed to determine TV (total vanadium) content and calculate vanadium leaching rate. The leachate is subjected to P-removal treatment and Ca-removal treatment to make P satisfy TV/P≥IOOO and [Ca +]<0.05g/L, and then extraction is carried out.
(3) Extraction 10-20volume% kerosene solution of P204 is saponified by ammonia. The saponified organic phase is mixed with leachate to perform one-stage extraction at room temperature for 5min under stirring. The mixture is kept standing for the phase separation, and the organic substances remained in the aqueous phase are removed The weight ratio of NH3 /Mn in the aqueous phase solution is controlled to be 0.6-50 : 1 by controlling saponification ratio and the extraction phase ratio .
Note: P204 represents bis-(2-ethylhexyl) phosphate.
(4) Vanadium precipitation and V2O5 production by calcination
A small amount of sulfuric acid is used to regulate pH of the oil-removed aqueous phase to 1.5-2.5, then the resultant is heated to above 90 °C, held for 60-120min, and filtered, the precipitate is washed with tap water containing [Na+K]<0. lg/L for 3 times and water used for each time is 30ml, and the washing wastewater is combined with supernatant of vanadium precipitation to give wastewater of vanadium precipitation. Ammonium polyvanadate precipitate is dried, calcined and melted at 500°C-800°C to give V2O5, and the composition of the obtained V2O5 is analyzed.
(5) Wastewater treatment
Lime milk with low water content is prepared, added into the wastewater of vanadium precipitation to regulate pH of the solution to 9.0-10.0, and filtered. The filtrate is regulated to pH of 5-7 with diluted sulfuric acid to give circulation water as residue washing water for leaching in the next cycle. The neutralization residue obtained from the filtration is used as raw material for Mn recovery.
(6) Regeneration of the organic phase The organic phase containing cation such as Mn2+ (carried organic phase) is reused after being reverse extracted by 2.5M sulfuric acid solution. After 10 cycles, the organic phase is reverse extracted by 6N hydrochloric acid besides the 2.5M sulfuric acid solution, and washed by dilute sulfuric acid to remove the Cl", and then the extraction agent returns to the system. The solution resulted from the hydrochloric acid reverse extraction and the solution resulted from the sulfuric acid solution reverse extraction are separately treated. When Mn in the reverse-extraction solution accumulates to an extent, the reverse-extraction solution reacts with wastewater neutralization residue to recover Mn.
The above steps (2)-(6) are repeated, 50Og of roasted material is used for each cycle, liquid solid ratio for each cycle is 4:1, vanadium precipitation is not carried out after leaching in the first cycle, the leachate is used for slurry preparation of the second cycle to increase vanadium concentration of the leachate; then during leaching in each of the rest cycles, residue washing water of last cycle is used in step (2), and the insufficient part is supplemented by circulation water; circulation water is used for washing the residue, and the insufficient part is supplemented by clear water. 51 cycles are carried out, and no wastewater containing ammonia nitrogen is discharged from the system.
Except the first cycle without vanadium precipitation, the maximum value, the minimum value, and the average value of the recovery rate of the rest cycles are shown in Table 2; the maximum value, the minimum value, and the average value of the chemical compositions of the V2O5 product of the rest cycles are shown in Table 3, and the Table 3 also shows compositions of the metallurgical No. 98 and No. 99 of Chinese National Standard No. GB3283-87; and the maximum value, the minimum value, and the average value of the composition of the circulation water after circulation treatment for the rest cycles are shown in Table 4.
Table 2 Vanadium recovery rate in circulation process (%)
Figure imgf000013_0001
The recovery rates of the raw material pretreatment and ammonia removal by calcination in the above Table 2 are obtained from long term industrial production data.
Note: The procedure of raw material pretreatment includes pulverization of bulk coarse vanadium slag, grinding, and iron removal, and a small amount of vanadium is lost in this process.
It can be observed from Table 2 that the total average recovery rate Of V2O5 from vanadium slag reaches 84.20% by the inventive method.
Table 3 Main components of V2Os product obtained by the circulation process(%)
Figure imgf000013_0002
The table 3 shows that the vanadium product obtained by the process has good product quality, and meets Chinese National Standard No. GB3283-87.
Table 4 Main components of circulation water (g/L)
Figure imgf000013_0003
It can be observed from Table 4 that, after 51 cycles of wastewater circulation, various impurity ions have no enrichment tendency, and circulation and reuse of vanadium extraction wastewater in low cost is realized, wherein K and Na are mainly from vanadium slag raw material, and sum of K+Na in the circulation water is stabilized within 0.2-0.4g/L and does not increase after 51 cycles.
Example 2 Roasted material prepared according to step (1) is adopted, 51 cycles of wastewater circulation test are carried out according to the inventive method, and each cycle includes steps (2)-(6).
(1) Preparation of roasted material:
The vanadium slag shown in Table 5 is ground to less than 0.098mm.
Table 5 Main components of vanadium slag (%)
Figure imgf000014_0001
28kg of ground vanadium slag powder is mixed with 1.82kg of lime (ground to below 0. lmm) containing 98% CaO, and then the mixture is roasted at 950 °C in air for 60min, cooled, and ground to less than 0.18mm. (2) Leaching of roasted material
1250ml of water resulted from residue washing of last cycle (clear water is used for the first cycle) is added to 50Og of the ground roasted material to prepare slurry, 32-65wt% sulfuric acid solution is slowly added continuously while stirring, pH is controlled at 2.8-3.3 during leaching process while the slurry temperature is held at a temperature between room temperature-58°C, the reaction is carried out for 60min. The resultant is filtered to give leachate, the residue is washed with circulation water (clear water is used for the first cycle) for 6 times, water used for each time is 120ml, the washing water is combined for preparing slurry for leaching of next cycle, and the residue is dried and weighed to determine TV (total vanadium) content and calculate vanadium leaching rate. The leachate is subjected to P-removal treatment and Ca-removal treatment to make P satisfy TV/P≥IOOO and [Ca +]<0.05g/L, and then extraction is carried out. (3) Extraction
10-30volume% kerosene solution of P507 is saponified by ammonia. The saponified organic phase is mixed with leachate to perform one-stage extraction at room temperature for 5min under stirring. The mixture is kept standing for the phase separation, and the organic substances remained in the aqueous phase are removed The weight ratio of NH3 /Mn in the aqueous phase solution is controlled to be 10-200 : 1 by controlling saponification ratio and the extraction phase ratio. Note: P507 represents mono-(2-ethylhexyl) 2-ethylhexylphosphonate. (4) Vanadium precipitation and V2O5 production by calcination
A small amount of sulfuric acid is used to regulate pH of the aqueous phase solution to 1.5-2.5, then the resultant is heated to above 90 °C, held for 60-120min, and filtered, the precipitate is washed with tap water containing [Na+K]<0.1g/L for 3 times and water used for each time is 30ml, and the washing wastewater is combined with supernatant of vanadium precipitation to give wastewater of vanadium precipitation.
Ammonium polyvanadate precipitate is dried, calcined and melted at 500°C-800°C to give V2O5, and the composition of the obtained V2O5 is analyzed. (5) Wastewater treatment
Lime milk with low water content is prepared, added into the wastewater of vanadium precipitation to regulate pH of the solution to 9.5-10.00, and filtered. The filtrate is regulated to pH of 5-7 with diluted sulfuric acid to give circulation water as residue washing water for leaching in the next cycle. The neutralization residue obtained from the filtration is used as raw material for Mn recovery. (6) Regeneration of the organic phase
The organic phase containing cation such as Mn2+ is reused after being reverse extracted by 2.5M sulfuric acid solution. After 10 cycles, the organic phase is reverse extracted by 6N hydrochloric acid besides the 2.5M sulfuric acid solution, and washed by dilute sulfuric acid to remove the Cl", and then the extraction agent returns to the system. The solution resulted from the hydrochloric acid reverse extraction and the solution resulted from the sulfuric acid solution reverse extraction are separately treated. When Mn in the reverse-extraction solution accumulates to an extent, the reverse-extraction solution reacts with wastewater neutralization residue to recover Mn. The above steps (2)-(6) are repeated, 50Og of roasted material is used for each cycle, liquid solid ratio for each cycle is 2.5:1, vanadium precipitation is not carried out after leaching in the first cycle, the leachate is used for slurry preparation of the second cycle to increase vanadium concentration of the leachate; then during leaching in each of the rest cycles, residue washing water of last cycle is used in step (2), and the insufficient part is supplemented by circulation water; circulation water is used for washing the residue, and the insufficient part is supplemented by clear water. 51 cycles are carried out, and no wastewater containing ammonia nitrogen is discharged from the system. Except the first cycle without vanadium precipitation, the maximum value, the minimum value, and the average value of the recovery rate of the rest cycles are shown in Table 6; the maximum value, the minimum value, and the average value of the chemical compositions of the V2O5 product of the rest cycles are shown in Table 7; and the maximum value, the minimum value, and the average value of the composition of the circulation water after circulation treatment for the rest cycles are shown in Table 8.
Table 6 Vanadium recovery rate in circulation process (%)
Figure imgf000016_0001
The recovery rates of the raw material pretreatment and ammonia removal by calcination in the above Table 6 are obtained from long term industrial production data. It can be observed from Table 6 that the total average recovery rate of V2O5 from vanadium slag reaches 84.48% by the inventive method.
Table 7 Main components of V2O5 product obtained by the circulation process(%)
Figure imgf000016_0002
The table 7 shows that the vanadium product obtained by the process has good product quality, and meets Chinese National Standard No. GB3283-87.
Table 8 Main components of circulation water (g/L)
Figure imgf000016_0003
It can be observed from Table 8 that, after 51 cycles of wastewater circulation, various impurity ions have no enrichment tendency, and circulation and reuse of vanadium extraction wastewater in low cost is realized, wherein K and Na are mainly from vanadium slag raw material, and sum of K+Na in the circulation water is stabilized within 0.1-0.4g/L and does not increase after 51 cycles.
Example 3
Roasted material prepared according to step (1) is adopted, 100 cycles of wastewater circulation test are carried out according to the inventive method, and each cycle includes steps (2)-(6). (1) Preparation of roasted material: 110kg of vanadium slag with composition shown in Table 9 is ground to less than 0.098mm, the ground vanadium slag powder is mixed with 7.7kg of lime (ground to below 0.1mm) containing 98% CaO, and then the mixture is roasted at 920 °C in air for 150min, cooled, and ground to less than 0.18mm.
Table 9 Main components of vanadium slag (%)
Figure imgf000017_0001
(2) Leaching of roasted material
2,000ml of water resulted from residue washing of last cycle (clear water is used for the first cycle) is added to 1,00Og of ground roasted material to prepare slurry, 50-75wt% sulfuric acid solution is slowly added continuously while stirring, pH is controlled at 2.8-3.3 during leaching process while the slurry temperature is held at a temperature between room temperature and 58°C, the reaction is carried out for 60min. The resultant is filtered to give leachate, the residue is washed with circulation water (clear water is used for the first cycle) for 5-6 times, water used for each time is 250ml, the washing water is combined for preparing slurry for leaching of next cycle, and the residue is dried and weighed to determine TV (total vanadium) content and calculate vanadium leaching rate. The leachate is subjected to P-removal treatment and Ca-removal treatment to make P satisfy TV/P≥IOOO and [Ca +]<0.05g/L, and then extraction is carried out.
(3) Extraction
10-30volume% kerosene solution of Cyanex is saponified by ammonia. The saponified organic phase is mixed with leachate to perform one-stage extraction at room temperature for 5min under stirring. The mixture is kept standing for the phase separation, and the organic substances remained in the aqueous phase are removed The weight ratio of NH3 /Mn in the aqueous phase solution is controlled to be 20-2000 : 1 by controlling saponification ratio and the extraction phase ratio. Note: Cyanex represents bis-(2,4,4-trimethylpentyl) hypophosphorous acid.
(4) Vanadium precipitation and V2O5 production by calcination
A small amount of sulfuric acid is used to regulate pH of the aqueous phase solution to 1.5-2.5, then the resultant is heated to above 90 °C, held for 60-120min, and filtered, the precipitate is washed with tap water containing [Na+K]<0.1g/L for 3 times and water used for each time is 30ml, and the washing wastewater is combined with supernatant of vanadium precipitation to give wastewater of vanadium precipitation. Ammonium polyvanadate precipitate is dried, calcined and melted at 500°C-800°C to give V2O5, and the composition of the obtained V2O5 is analyzed.
(5) Wastewater treatment
Lime milk with low water content is prepared, added into the wastewater of vanadium precipitation to regulate pH of the solution to 9.5-11, and filtered. The filtrate is regulated to pH of 5-7 with diluted sulfuric acid to give circulation water as residue washing water for leaching in the next cycle. The neutralization residue obtained from the filtration is used as raw material for Mn recovery.
(6) Regeneration of the organic phase
The organic phase containing cation such as Mn2+ is reused after being reverse extracted by 2.5M sulfuric acid solution. After 10 cycles, the organic phase is reverse extracted by 6N hydrochloric acid besides the 2.5M sulfuric acid solution, and washed by dilute sulfuric acid to remove the Cl", and then the extraction agent returns to the system. The solution resulted from the hydrochloric acid reverse extraction and the solution resulted from the sulfuric acid solution reverse extraction are separately treated. When Mn in the reverse-extraction solution accumulates to an extent, the reverse-extraction solution reacts with wastewater neutralization residue to recover Mn.
The above steps (2)-(6) are repeated, lOOOg of roasted material is used for each cycle, liquid solid ratio for each cycle is 2:1, vanadium precipitation is not carried out after leaching in the first cycle, the leachate is used for slurry preparation of the second cycle to increase vanadium concentration of the leachate; then during leaching in each of the rest cycles, residue washing water of last cycle is used in step (2), and the insufficient part is supplemented by circulation water; circulation water is used for washing the residue, and the insufficient part is supplemented by clear water. 51 cycles are carried out, and no wastewater containing ammonia nitrogen is discharged from the system. Except the first cycle without vanadium precipitation, the maximum value, the minimum value, and the average value of the recovery rate of the rest cycles are shown in Table 10; the maximum value, the minimum value, and the average value of the chemical compositions of the V2O 5 product of the rest cycles are shown in Table 11; and the maximum value, the minimum value, and the average value of the composition of the circulation water after circulation treatment for the rest cycles are shown in Table 12.
Table 10 Vanadium recovery rate in circulation process (%)
Figure imgf000018_0001
Figure imgf000019_0001
The recovery rates of the raw material pretreatment and ammonia removal by calcination in the above Table are obtained from long term industrial production data.
It can be observed from Table 10 that the total average recovery rate of V2O5 from vanadium slag reaches 83.89% by the inventive method.
Table 11 Main components of V2Os product obtained by the circulation process(%)
Figure imgf000019_0002
The table 11 shows that the vanadium product obtained by the process has good product quality, and meets Chinese National Standard No. GB3283-87.
Table 12 Main components of circulation water (g/L)
Figure imgf000019_0003
It can be observed from Table 12 that, after 51 cycles of wastewater circulation, various impurity ions have no enrichment tendency, and circulation and reuse of vanadium extraction wastewater in low cost is realized, wherein K and Na are mainly from vanadium slag raw material, and sum of K+Na in the circulation water is stabilized within 0.2-0.4g/L and does not increase after 51 cycles.

Claims

Claims
1. A production method of vanadium oxide using extraction, comprising the following steps: a. mixing vanadium -containing material with additive to give mixed material, wherein the additive is CaO or limestone, and the amount of the additive makes the CaOAV2O5 weight ratio in the mixed material be 0.5-1.4 : 1; b. roasting the mixed material at 860°C-950°C in oxidizing atmosphere for 60-240min to give roasted material; c. adding water into the roasted material to give slurry, stirring, and slowly adding sulfuric acid solution to leach the slurry while controlling pH at 2.5-3.5; d. removing residue after leaching to give leachate, subjecting the leachate to P -removal treatment and Ca-removal treatment to make weight ratio of total V to P in the leachate >1000 and [Ca +] in the leachate <0.05g/L, washing the residue with circulation water to give washing water useful for preparing slurry in next leaching; e. subjecting the P-removed and Ca-removed leachate to extraction by using an extraction agent which has been saponified by ammonia, and then separating the aqueous phase and organic phase, the weight ratio of NH3AMn in the aqueous phase being controlled within 0.6-2000 : 1, the extraction agent being capable of extracting Mn2+ ,Mg2+ and Fe3+ at pH of2-5; f. regulating pH of the aqueous phase with sulfuric acid to 1.5-2.5, heating to a temperature between 90 °C and boiling temperature, holding at the temperature for 30-120min, filtering, and washing and drying the precipitate to give ammonium polyvanadate, subjecting the ammonium polyvanadate to calcination to remove ammonia to give V2O5, or reducing the ammonium polyvanadate to give V2O3; and g. removing P, Mn, and Mg impurities in the wastewater generated in step f to make Mn2+ and Mg2+concentrations lower than 5g/L respectively, and P concentration lower than 0.005g/L to give circulation water, and returning the circulation water to step c for preparing slurry and to step d for washing the residue, wherein the solid raw material used in the above steps has total alkali metal amount not more than 0.3wt%, and total amount of Cl and NO3 not more than 0.1 wt%; and the liquid raw material used in the above steps has total alkali metal amount not more than 0.1 g/L, and total amount of Cl and NO3 not more than 0.1 g/L.
2. The production method of vanadium oxide according to claim 1, wherein the mixed material obtained in step a has particle size of 0.1mm or less.
3. The production method of vanadium oxide according to claim 1, wherein in step c, the roasted material is cooled and pulverized to 0.18mm or less before being prepared into slurry.
4. The production method of vanadium oxide according to claim 1, wherein in step c, the concentration of the sulfuric acid solution is 10-75wt%, the leaching temperature is
10-58°C, and the leaching time is 30-90min.
5. The production method of vanadium oxide according to claim 1, wherein the circulation water obtained in step g is returned to step d to wash the residue to obtain washing water, the washing water is returned to step c for leaching and preparing slurry, if the washing water is not sufficient, the circulation water is adopted for supplement.
6. The production method of vanadium oxide according to claim 1 or 5, wherein in step c, the weight ratio of total amount of water for slurry preparation to the roasted material is 1.5-4 : 1.
7. The production method of vanadium oxide according to claim 1, wherein during washing of the residue, the washing times are 5-7, and the amount of water used for each washing is 20-35wt% of the residue on dry basis.
8. The production method of vanadium oxide according to claim 1, wherein in step e, the extraction agent is at least one of bis-(2-ethylhexyl) phosphate, mono-(2-ethylhexyl) 2-ethylhexylphosphonate, and bis-(2,4,4-trimethylpentyl) hypophosphorous acid..
9. The production method of vanadium oxide according to claim 1, wherein in step e, the extraction agent is previously saponified by ammonia such that the raffinate has pH of
2.8-3.8.
10. The production method of vanadium oxide according to claim 1, wherein in step e, the conditions of the extraction make weight ratio of NH3/Mn in the aqueous phase within 2-10 : 1.
11. The production method of vanadium oxide according to claim 1, wherein in step e, the organic phase is reused after being regenerated by 2-3M sulfuric acid solution and then being saponified by ammonia.
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