CN114350965B - Method for extracting vanadium, manganese and recycling wastewater from vanadium slag calcified roasting clinker - Google Patents
Method for extracting vanadium, manganese and recycling wastewater from vanadium slag calcified roasting clinker Download PDFInfo
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- 229910052720 vanadium Inorganic materials 0.000 title claims abstract description 221
- LEONUFNNVUYDNQ-UHFFFAOYSA-N vanadium atom Chemical compound [V] LEONUFNNVUYDNQ-UHFFFAOYSA-N 0.000 title claims abstract description 220
- PWHULOQIROXLJO-UHFFFAOYSA-N Manganese Chemical compound [Mn] PWHULOQIROXLJO-UHFFFAOYSA-N 0.000 title claims abstract description 100
- 229910052748 manganese Inorganic materials 0.000 title claims abstract description 89
- 239000011572 manganese Substances 0.000 title claims abstract description 89
- 239000002351 wastewater Substances 0.000 title claims abstract description 82
- 239000002893 slag Substances 0.000 title claims abstract description 78
- 238000000034 method Methods 0.000 title claims abstract description 60
- 238000004064 recycling Methods 0.000 title claims abstract description 28
- 238000005406 washing Methods 0.000 claims abstract description 122
- 239000007788 liquid Substances 0.000 claims abstract description 95
- 238000002386 leaching Methods 0.000 claims abstract description 93
- 238000000605 extraction Methods 0.000 claims abstract description 59
- 239000000706 filtrate Substances 0.000 claims abstract description 55
- 239000000243 solution Substances 0.000 claims abstract description 52
- 239000012535 impurity Substances 0.000 claims abstract description 35
- 239000003795 chemical substances by application Substances 0.000 claims abstract description 25
- 238000002156 mixing Methods 0.000 claims abstract description 14
- 230000003472 neutralizing effect Effects 0.000 claims abstract description 10
- 230000001376 precipitating effect Effects 0.000 claims abstract description 9
- 239000010413 mother solution Substances 0.000 claims abstract 4
- 239000012452 mother liquor Substances 0.000 claims description 55
- 239000007787 solid Substances 0.000 claims description 37
- 229910052698 phosphorus Inorganic materials 0.000 claims description 27
- 239000000047 product Substances 0.000 claims description 27
- GNTDGMZSJNCJKK-UHFFFAOYSA-N divanadium pentaoxide Chemical compound O=[V](=O)O[V](=O)=O GNTDGMZSJNCJKK-UHFFFAOYSA-N 0.000 claims description 24
- 238000005868 electrolysis reaction Methods 0.000 claims description 21
- 239000000654 additive Substances 0.000 claims description 15
- 230000000996 additive effect Effects 0.000 claims description 15
- QGZKDVFQNNGYKY-UHFFFAOYSA-O Ammonium Chemical compound [NH4+] QGZKDVFQNNGYKY-UHFFFAOYSA-O 0.000 claims description 14
- 238000003756 stirring Methods 0.000 claims description 11
- 238000001354 calcination Methods 0.000 claims description 8
- 229910052751 metal Inorganic materials 0.000 claims description 8
- 239000002184 metal Substances 0.000 claims description 8
- 150000003863 ammonium salts Chemical class 0.000 claims description 3
- 238000011084 recovery Methods 0.000 abstract description 4
- 238000009854 hydrometallurgy Methods 0.000 abstract description 2
- OAICVXFJPJFONN-UHFFFAOYSA-N Phosphorus Chemical compound [P] OAICVXFJPJFONN-UHFFFAOYSA-N 0.000 description 24
- 239000011574 phosphorus Substances 0.000 description 24
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 description 14
- JPJALAQPGMAKDF-UHFFFAOYSA-N selenium dioxide Chemical group O=[Se]=O JPJALAQPGMAKDF-UHFFFAOYSA-N 0.000 description 14
- 230000000052 comparative effect Effects 0.000 description 13
- 238000001556 precipitation Methods 0.000 description 11
- 239000002253 acid Substances 0.000 description 7
- 230000001276 controlling effect Effects 0.000 description 6
- 238000004537 pulping Methods 0.000 description 5
- 238000000926 separation method Methods 0.000 description 5
- 230000002378 acidificating effect Effects 0.000 description 4
- 230000002308 calcification Effects 0.000 description 4
- 239000012670 alkaline solution Substances 0.000 description 3
- BFNBIHQBYMNNAN-UHFFFAOYSA-N ammonium sulfate Chemical compound N.N.OS(O)(=O)=O BFNBIHQBYMNNAN-UHFFFAOYSA-N 0.000 description 3
- 229910052921 ammonium sulfate Inorganic materials 0.000 description 3
- 235000011130 ammonium sulphate Nutrition 0.000 description 3
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 3
- 235000008733 Citrus aurantifolia Nutrition 0.000 description 2
- 235000011941 Tilia x europaea Nutrition 0.000 description 2
- XHCLAFWTIXFWPH-UHFFFAOYSA-N [O-2].[O-2].[O-2].[O-2].[O-2].[V+5].[V+5] Chemical compound [O-2].[O-2].[O-2].[O-2].[O-2].[V+5].[V+5] XHCLAFWTIXFWPH-UHFFFAOYSA-N 0.000 description 2
- 159000000007 calcium salts Chemical class 0.000 description 2
- OSGAYBCDTDRGGQ-UHFFFAOYSA-L calcium sulfate Chemical compound [Ca+2].[O-]S([O-])(=O)=O OSGAYBCDTDRGGQ-UHFFFAOYSA-L 0.000 description 2
- 238000006243 chemical reaction Methods 0.000 description 2
- 239000003792 electrolyte Substances 0.000 description 2
- 239000004571 lime Substances 0.000 description 2
- WPBNNNQJVZRUHP-UHFFFAOYSA-L manganese(2+);methyl n-[[2-(methoxycarbonylcarbamothioylamino)phenyl]carbamothioyl]carbamate;n-[2-(sulfidocarbothioylamino)ethyl]carbamodithioate Chemical compound [Mn+2].[S-]C(=S)NCCNC([S-])=S.COC(=O)NC(=S)NC1=CC=CC=C1NC(=S)NC(=O)OC WPBNNNQJVZRUHP-UHFFFAOYSA-L 0.000 description 2
- 239000008267 milk Substances 0.000 description 2
- 210000004080 milk Anatomy 0.000 description 2
- 235000013336 milk Nutrition 0.000 description 2
- 238000006386 neutralization reaction Methods 0.000 description 2
- 238000006722 reduction reaction Methods 0.000 description 2
- 230000001105 regulatory effect Effects 0.000 description 2
- 229910001935 vanadium oxide Inorganic materials 0.000 description 2
- BHPQYMZQTOCNFJ-UHFFFAOYSA-N Calcium cation Chemical compound [Ca+2] BHPQYMZQTOCNFJ-UHFFFAOYSA-N 0.000 description 1
- 239000007864 aqueous solution Substances 0.000 description 1
- 230000015572 biosynthetic process Effects 0.000 description 1
- 229910001424 calcium ion Inorganic materials 0.000 description 1
- 238000009792 diffusion process Methods 0.000 description 1
- 238000001035 drying Methods 0.000 description 1
- 230000005611 electricity Effects 0.000 description 1
- 238000005265 energy consumption Methods 0.000 description 1
- 238000001704 evaporation Methods 0.000 description 1
- 239000010440 gypsum Substances 0.000 description 1
- 229910052602 gypsum Inorganic materials 0.000 description 1
- 239000011964 heteropoly acid Substances 0.000 description 1
- 229910052739 hydrogen Inorganic materials 0.000 description 1
- 239000001257 hydrogen Substances 0.000 description 1
- -1 hydrogen ions Chemical class 0.000 description 1
- 239000010842 industrial wastewater Substances 0.000 description 1
- 238000004519 manufacturing process Methods 0.000 description 1
- 150000002739 metals Chemical class 0.000 description 1
- JKJKPRIBNYTIFH-UHFFFAOYSA-N phosphanylidynevanadium Chemical compound [V]#P JKJKPRIBNYTIFH-UHFFFAOYSA-N 0.000 description 1
- 230000010287 polarization Effects 0.000 description 1
- 238000004886 process control Methods 0.000 description 1
- 230000001737 promoting effect Effects 0.000 description 1
- 239000002994 raw material Substances 0.000 description 1
- 239000002002 slurry Substances 0.000 description 1
- 239000006104 solid solution Substances 0.000 description 1
- 239000002699 waste material Substances 0.000 description 1
Classifications
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
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- Inorganic Compounds Of Heavy Metals (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention relates to the technical field of hydrometallurgy, and discloses a method for extracting vanadium, manganese and recycling wastewater from vanadium slag calcified roasting clinker. The method comprises the following steps: mixing clinker and a first mother solution, adding a second mother solution, and adding a leaching agent for leaching to obtain residues and leaching liquid; washing residues by using a first washing liquid, a second washing liquid and a third washing liquid to obtain a first washing filtrate, a second washing filtrate and a third washing filtrate; precipitating vanadium from the first washing filtrate and the leaching solution to obtain vanadium extraction wastewater; removing impurities from part of vanadium extraction wastewater, neutralizing to obtain an impurity removing solution, and carrying out electrolytic treatment on part of the impurity removing solution to obtain an electrolytic anode solution; returning the second washing filtrate to use, and returning the third washing filtrate to use; and returning the residual vanadium extraction wastewater to use, returning the residual impurity removal liquid to use, and returning the electrolytic anolyte to use. The method realizes the effective recycling of the electrolytic anolyte and the washing wastewater, and the efficient leaching of vanadium and the low-cost recovery of manganese in calcified clinker.
Description
Technical Field
The invention relates to the technical field of hydrometallurgy, in particular to a method for extracting vanadium, manganese and recycling wastewater from vanadium slag calcified roasting clinker.
Background
In the calcified vanadium extraction process, the converter vanadium slag is crushed and ball milled and then added with a certain amount of calcium salt for roasting, low-valence vanadium in the vanadium slag is oxidized into pentavalent vanadium, and the vanadium oxide product is obtained through leaching, precipitation, drying and reduction. As the raw material vanadium slag contains not only vanadium but also a large amount of manganese, the manganese coexists with vanadium in vanadium precipitation qualified liquid after roasting, leaching and other processes, and then is separated from vanadium in precipitation process, wherein vanadium exists in product APV, and manganese exists in vanadium extraction wastewater. At present, the treatment modes adopted by the vanadium extraction wastewater mainly comprise a lime milk neutralization method and a reduction neutralization-evaporation concentration method, and the problems of large slag quantity, difficult recycling of slag, high treatment cost and the like generally exist.
Aiming at improving the vanadium leaching rate of calcified clinker, patent CN 109338103B discloses a method for leaching vanadium by countercurrent acid leaching of calcified clinker, which realizes the gradual leaching of vanadium by controlling the pH value of the three-stage leaching process or the leaching end point to be 2.8-3.5, 1.8-3.5 and 0.7-1.5 respectively; and then the leaching washing filtrate of each stage is returned to the previous stage to realize the utilization of the low-concentration vanadium-containing solution. The method can effectively improve the leaching rate of vanadium (reaching more than 90 percent), but has more process control points (the leaching pH of each stage, the filtrate pH and the vanadium concentration need to be controlled), and the three-stage leaching needs to be controlled to be between 0.7 and 1.5, and the pH of the final leaching liquid needs to be between 1.4 and 2.2.
Aiming at vanadium precipitation wastewater generated by a calcification roasting process, patent CN102838233B adopts lime milk to adjust the acidic vanadium precipitation wastewater to be alkaline, and then solid-liquid separation is carried out to obtain alkaline solution and gypsum slag; the pH value is regulated to 3-7 after decalcification of the alkaline solution, and then the alkaline solution is returned to the leaching process for recycling. The method realizes the recycling of wastewater, can effectively control the concentration of calcium ions in the solution returned to the leaching process, avoids the formation of calcium sulfate precipitation, but does not effectively recycle a large amount of manganese resources in the wastewater, and has low ore utilization rate. Patent CN 104058523B proposes a treatment method for industrial wastewater from vanadium oxide production, which recovers metal manganese products by means of impurity removal-electrolysis, and does not relate to a recycling method for anolyte obtained after electrolysis. Patent CN 105219969B provides a method for extracting manganese metal by utilizing vanadium precipitation wastewater and vanadium extraction tailings, leaching the vanadium precipitation wastewater and the vanadium extraction tailings, removing impurities from the obtained leaching solution, electrolyzing to recover manganese metal, and returning part of anolyte generated by electrolysis to the vanadium extraction tailings and part to the clinker leaching process, but does not give a detailed returned clinker leaching utilization method. Meanwhile, the concentration of manganese in the vanadium extraction wastewater is low, about 8-15g/L, and when the electrolytic method is adopted to extract metal manganese, the problems of low current efficiency in the electrolytic process, increased control difficulty in the electrolytic process and the like exist.
Disclosure of Invention
The invention aims to solve the problems of low manganese resource recycling rate, high electricity consumption cost in the manganese extraction process, high low-pH solution recycling difficulty and the like in the prior art, and provides a method for extracting vanadium, manganese and recycling wastewater by using vanadium slag calcified roasting clinker.
In order to achieve the above purpose, the invention provides a method for extracting vanadium, manganese and recycling waste water from vanadium slag calcified roasting clinker, which comprises the following steps:
(1) Mixing the vanadium slag calcified roasting clinker with the first mother liquor under stirring, adding the second mother liquor under stirring, adding a leaching agent, and leaching under the condition that the pH value is 2.6-3.5 to obtain residues and leaching liquid;
(2) Washing residues obtained in the step (1) sequentially by using a first washing liquid, a second washing liquid and a third washing liquid to obtain a first washing filtrate, a second washing filtrate, a third washing filtrate and vanadium extraction tailings respectively;
(3) Mixing the first washing filtrate with the leaching solution, adding ammonium salt for precipitating vanadium to obtain ammonium polyvanadate and vanadium extraction wastewater, and calcining the ammonium polyvanadate to obtain a vanadium pentoxide product;
(4) Removing impurities from part of vanadium extraction wastewater, neutralizing to obtain an impurity-removing solution, and then adding an electrolysis additive into part of the impurity-removing solution for electrolysis treatment to obtain an electrolysis anolyte and a metal manganese product;
(5) Returning the second washing filtrate to the step (1) to be used as a first mother liquor, and returning the third washing filtrate to the step (1) to be used as a second mother liquor;
(6) And (3) returning the residual vanadium extraction wastewater to the step (2), returning the residual impurity removal liquid to the step (2), and returning the electrolytic anolyte to the step (2) and/or the step (1).
Preferably, in the step (1), the vanadium slag calcified roasting clinker contains 12-25 wt% of V 2 O 5 5 to 10 weight percent of MnO and less than or equal to 0.1 weight percent of P.
Preferably, in the step (1), the liquid-solid ratio of the first pulping mother liquor to the vanadium slag calcified roasting clinker is 0.6-2mL/g;
preferably, the liquid-solid ratio of the second pulping mother liquor to the vanadium slag calcified roasting clinker is 0.8-1.5mL/g;
preferably, the liquid-solid ratio of the leaching agent to the vanadium slag calcified roasting clinker is 0.5-1.5mL/g;
preferably, the ratio of the total volume of the first mother liquor, the second mother liquor and the leaching agent to the weight of the vanadium slag calcified roasting clinker is 2-4mL/g.
Preferably, in step (1), the leaching time is 40-60min.
Preferably, in step (1), the leach solution contains 15-40g/L manganese.
Preferably, in the step (2), the liquid-solid ratio of the first washing liquid to the vanadium slag calcified roasting clinker is 0.2-0.5mL/g.
Preferably, in the step (2), the liquid-solid ratio of the second washing liquid to the vanadium slag calcified roasting clinker is 0.6-2mL/g.
Preferably, in the step (2), the liquid-solid ratio of the third washing liquid to the vanadium slag calcified roasting clinker is 0.8-1.5mL/g.
Preferably, in the step (4), the concentration of manganese in the cathode region during the electrolytic treatment is controlled to be 8-15g/L, and the current density during the electrolytic treatment is controlled to be 260-350A/m 2 。
Preferably, when the pH value of the residual vanadium extraction wastewater is=1.2-4 and the P content is less than or equal to 0.08g/L, the residual vanadium extraction wastewater can be directly returned to the step (2) for use; when the vanadium extraction wastewater does not meet the requirements, the wastewater needs to be adjusted to be within the required range, and then the wastewater returns to the step (2) for use.
The method adopts the modes of clinker pulping, acid leaching and three times of washing to realize the effective recycling of electrolytic anolyte and washing wastewater, and the efficient leaching of vanadium and the low-cost recovery of manganese in calcified clinker. The method has the characteristics of simple and easy use process, wide application range and low cost, and has high social and economic benefits.
Detailed Description
The following describes specific embodiments of the present invention in detail. It should be understood that the detailed description and specific examples, while indicating and illustrating the invention, are not intended to limit the invention.
The endpoints and any values of the ranges disclosed herein are not limited to the precise range or value, and are understood to encompass values approaching those ranges or values. For numerical ranges, one or more new numerical ranges may be found between the endpoints of each range, between the endpoint of each range and the individual point value, and between the individual point value, in combination with each other, and are to be considered as specifically disclosed herein.
The invention provides a method for extracting vanadium, manganese and recycling wastewater from vanadium slag calcified roasting clinker, which comprises the following steps:
(1) Mixing the vanadium slag calcified roasting clinker with the first mother liquor under stirring, adding the second mother liquor under stirring, adding a leaching agent, and leaching under the condition that the pH value is 2.6-3.5 to obtain residues and leaching liquid;
(2) Washing residues obtained in the step (1) sequentially by using a first washing liquid, a second washing liquid and a third washing liquid to obtain a first washing filtrate, a second washing filtrate, a third washing filtrate and vanadium extraction tailings respectively;
(3) Mixing the first washing filtrate with the leaching solution, adding ammonium salt for precipitating vanadium to obtain ammonium polyvanadate and vanadium extraction wastewater, and calcining the ammonium polyvanadate to obtain a vanadium pentoxide product;
(4) Removing impurities from part of vanadium extraction wastewater, neutralizing to obtain an impurity-removing solution, and then adding an electrolysis additive into part of the impurity-removing solution for electrolysis treatment to obtain an electrolysis anolyte and a metal manganese product;
(5) Returning the second washing filtrate to the step (1) to be used as a first mother liquor, and returning the third washing filtrate to the step (1) to be used as a second mother liquor;
(6) And (3) returning the residual vanadium extraction wastewater to the step (2), returning the residual impurity removal liquid to the step (2), and returning the electrolytic anolyte to the step (2) and/or the step (1).
In the invention, the vanadium slag calcified roasting clinker is obtained by mixing and roasting the vanadium slag and calcium salt of a converter, and preferably, the vanadium slag calcified roasting clinker contains 12 to 25 weight percentV 2 O 5 5 to 10 weight percent of MnO and less than or equal to 0.1 weight percent of P.
According to the invention, the first mother liquor and the clinker are stirred and mixed, and then the second mother liquor is slowly added in a stirring state to improve the liquid-solid ratio in the leaching process, meanwhile, the phenomenon of vanadium precipitation caused by difficult diffusion of hydrogen ions in the leaching process due to thick slurry can be avoided, and then the acidity regulator is added for leaching under an acidic condition, so that the effective leaching of vanadium in the calcified clinker is realized.
In a preferred embodiment, in step (1), the liquid-to-solid ratio of the first mother liquor to the vanadium slag calcified calcined clinker is from 0.6 to 2mL/g. Specifically, it may be 0.6mL/g, 0.7mL/g, 0.8mL/g, 0.9mL/g, 1mL/g, 1.1mL/g, 1.2mL/g, 1.3mL/g, 1.4mL/g, 1.5mL/g, 1.6mL/g, 1.7mL/g, 1.8mL/g, 1.9mL/g or 2mL/g.
In a preferred embodiment, the liquid to solid ratio of the second mother liquor to the vanadium slag calcified-roasting clinker is 0.8-1.5mL/g. Specifically, it may be 0.8mL/g, 0.9mL/g, 1mL/g, 1.1mL/g, 1.2mL/g, 1.3mL/g, 1.4mL/g or 1.5mL/g.
In a preferred embodiment, in step (1), the total time for mixing the vanadium slag calcified roasting clinker with the first mother liquor and the second mother liquor is 5-35min.
In a preferred embodiment, in step (1), the leach solution contains 15-40g/L manganese.
In a preferred embodiment, in step (1), the first mother liquor is a manganese-containing solution having a ph=1.5-8, p.ltoreq.0.08 g/L.
In a preferred embodiment, in step (1), the second mother liquor is a manganese-containing solution having P.ltoreq.0.08 g/L.
It is further preferred that in step (1) the second mother liquor is slowly added and the pH of the system during addition is kept between 2.6 and 3.5.
Further preferably, the first mother liquor and the second mother liquor may be vanadium extraction wastewater generated in a conventional calcification roasting-acid leaching vanadium extraction process.
According to the invention, the vanadium extraction wastewater generated in the conventional calcification roasting-acid leaching vanadium extraction process is used as the first mother liquor and the second mother liquor, so that the recycling of the wastewater can be realized, the manganese element contained in the wastewater can be utilized, the manganese content in the solution in the electrolysis process can be increased, the manganese concentration in the cathode manganese separation area in the electrolysis process can be further increased, the concentration polarization can be reduced, and the electrolysis power consumption can be further reduced.
In the invention, the pH value of the system in the leaching process can be controlled to be 2.6-3.5 by adding sulfuric acid in the leaching process. In a preferred embodiment, the leaching agent may be water or an acidic aqueous solution or an acidic anolyte produced after electrolysis.
Further preferably, the liquid-solid ratio of the leaching agent to the vanadium slag calcified-roasting clinker is 0.5-1.5mL/g. Specifically, it may be 0.5mL/g, 0.6mL/g, 0.7mL/g, 0.8mL/g, 0.9mL/g, 1mL/g, 1.1mL/g, 1.2mL/g, 1.3mL/g, 1.4mL/g or 1.5mL/g.
In a preferred embodiment, the ratio of the total volume of the first mother liquor, the second mother liquor and the leaching agent to the weight of the vanadium slag calcified roasting clinker is 2-4mL/g.
Further preferably, in step (1), the leaching time is 40-60min. Specifically, it may be 40min, 45min, 50min, 55min or 60min.
In a preferred embodiment, the first washing liquid is a manganese-containing solution having a ph=1.5 to 8 and p.ltoreq.0.08 g/L.
In a preferred embodiment, the second washing liquid is a manganese-containing solution having a ph=1.5 to 8 and p.ltoreq.0.08 g/L.
In a preferred embodiment, the third washing liquid is a manganese-containing solution having a ph=0.6 to 1.5 and p.ltoreq.0.08 g/L.
In a preferred embodiment, in step (2), the liquid-to-solid ratio of the first washing liquid to the vanadium slag calcified calcined clinker is in the range of 0.2 to 0.5mL/g. Specifically, it may be 0.2mL/g, 0.3mL/g, 0.4mL/g or 0.5mL/g.
In a preferred embodiment, in step (2), the liquid-to-solid ratio of the second washing liquid to the vanadium slag calcified calcined clinker is in the range of 0.6-2mL/g. Specifically, it may be 0.6mL/g, 0.7mL/g, 0.8mL/g, 0.9mL/g, 1mL/g, 1.1mL/g, 1.2mL/g, 1.3mL/g, 1.4mL/g, 1.5mL/g, 1.6mL/g, 1.7mL/g, 1.8mL/g, 1.9mL/g or 2mL/g.
In a preferred embodiment, in step (2), the liquid-to-solid ratio of the third washing liquid to the vanadium slag calcified calcined clinker is in the range of 0.8 to 1.5mL/g. Specifically, it may be 0.8mL/g, 0.9mL/g, 1mL/g, 1.1mL/g, 1.2mL/g, 1.3mL/g, 1.4mL/g or 1.5mL/g.
In the invention, the second washing filtrate generated in the washing process can be recycled as the first mother liquor in the step (1), and the third washing filtrate can be recycled as the second mother liquor in the step (1), so that the consumption of the washing liquid cost and the discharge of the washing wastewater are reduced.
In a preferred embodiment, in the step (4), the P content in the impurity removing liquid obtained by impurity removing and neutralizing is less than or equal to 0.01g/L, and the pH value is 5.5-7.5. Preferably 6-6.5.
In a preferred embodiment, in step (4), the electrolysis additive is selenium dioxide. Further preferably, when the electrolytic additive is added to a part of the impurity removing liquid, the volume ratio of the weight of the electrolytic additive to the impurity removing liquid is 0.01 to 0.03g/L.
In a preferred embodiment, in step (4), the concentration of manganese in the cathode region during the electrolytic treatment is controlled to be 8-15g/L, and the current density during the electrolysis is controlled to be 260-350A/m 2 。
In the invention, in order to further reduce the wastewater discharge and the waste of resources, the residual vanadium extraction wastewater, the impurity removal liquid and the electrolytic anolyte can be returned to the step (2) and recycled in the step (1).
In the invention, the residual vanadium extraction wastewater can be returned to the step (2) to be used as the first washing liquid and/or the second washing liquid; when the pH value of the residual vanadium extraction wastewater is=1.2-4 and the P content is less than or equal to 0.08g/L, the residual vanadium extraction wastewater can be directly returned to the step (2) for use; when the vanadium extraction wastewater does not meet the requirements, the wastewater needs to be adjusted to be within the required range, and then the wastewater returns to the step (2) for use. Specifically, the vanadium precipitation wastewater can be adjusted to be within a required range by means of neutralization, impurity removal and the like, and then the wastewater can be returned to the step (2) for use.
In the present invention, the remaining impurity-removed liquid may be returned to step (2) to be used as the first washing liquid and/or the second washing liquid.
In the invention, the electrolyte anolyte obtained by electrolysis contains a large amount of sulfuric acid, and the pH value is usually 0.6-1.5, so that the electrolyte anolyte can be returned to the step (1) as a leaching agent for promoting leaching of vanadium, and/or returned to the step (2) as a third washing filtrate, vanadium in residues can be further dissolved, and the recovery rate of vanadium is further improved.
The method can effectively extract valuable metals vanadium and manganese in the vanadium slag calcified roasting clinker, and the wastewater produced in the whole process can be returned to the system for recycling. On the basis of keeping excellent extraction efficiency, zero wastewater discharge is realized, and the whole process is simple and easy to operate, has low cost and has high social and economic benefits.
The present invention will be described in detail by way of examples, but the scope of the present invention is not limited thereto.
Example 1
(1) Calcified roasting clinker (containing V) of vanadium slag 2 O 5 16.06%, mnO 8.64%, P0.05%) and a first mother liquor (containing 3.8g/L of vanadium, 12.2g/L of manganese and 0.03g/L, pH =2.8) are mixed for 5min, wherein the liquid-solid ratio of the first mother liquor to the vanadium slag calcified roasting clinker is 0.9mL/g, then a second mother liquor (containing 0.45g/L of vanadium, 10.5g/L of manganese and 0.06g/L, pH =2.6) is slowly added in 30min under stirring, the pH value of the system in the adding process is controlled to be 2.6-3.5, the liquid-solid ratio of the second mother liquor to the vanadium slag calcified roasting clinker is 1mL/g, then a leaching agent (containing 10.8g/L of manganese and 0.01g/L, pH =1.3) is slowly added, the leaching agent and the liquid-solid ratio of the vanadium slag calcified roasting clinker is 0.8mL/g under the condition of 2.8 by using sulfuric acid, and the leaching agent and the liquid-solid ratio of the vanadium slag calcified roasting clinker is controlled to be 2.6-3.5 g, and the leaching agent is separated to obtain a leaching solution (containing 21.03 g/L of vanadium and 37.03 g/L);
(2) Washing the residue obtained in the step (1) by sequentially using a first washing liquid (containing 18.9g/L of manganese and no more than 0.01g/L, pH =6.5), a second washing liquid (containing 0.18g/L of vanadium, 19.8g/L of manganese and 0.02g/L, pH =1.8 of phosphorus) and a third washing liquid (containing 10.8g/L of manganese and no more than 0.01g/L, pH =1.3) to respectively obtain a first washing filtrate, a second washing filtrate, a third washing filtrate and vanadium extraction tailings (TV=0.95%); wherein the liquid-solid ratio of the first washing liquid to the vanadium slag calcified roasting clinker is 0.3mL/g, the liquid-solid ratio of the second washing liquid to the vanadium slag calcified roasting clinker is 1mL/g, and the liquid-solid ratio of the third washing liquid to the vanadium slag calcified roasting clinker is 1mL/g;
(3) Mixing the first washing filtrate with the leaching solution, adding ammonium sulfate for precipitating vanadium to obtain ammonium polyvanadate and vanadium extraction wastewater (containing 0.25g/L of vanadium, 20.9g/L of manganese and 0.02g/L, pH =1.8 of phosphorus), and calcining the ammonium polyvanadate to obtain a vanadium pentoxide product;
(4) Removing impurities from part of vanadium extraction wastewater, neutralizing to obtain impurity-removing solution (vanadium not more than 0.01g/L, manganese not more than 20.1g/L, phosphorus not more than 0.01g/L, pH =6.0), adding electrolytic additive (selenium dioxide) into part of impurity-removing solution, performing electrolytic treatment, and controlling current density to 320A/m 2 The concentration of manganese in the cathode region is 8-12g/L, and electrolytic anolyte and manganese metal products are obtained; when the electrolytic additive is added into part of the impurity removing liquid, the volume ratio of the weight of the electrolytic additive to the impurity removing liquid is 0.02g/L;
(5) Returning the second washing filtrate to the step (1) to be used as a first mother liquor, and returning the third washing filtrate to the step (1) to be used as a second mother liquor;
(6) The residual vanadium extraction wastewater is directly returned to the step (2) to be used as a second washing liquid, the residual impurity removal liquid is returned to the step (2) to be used as a first washing liquid, the electrolytic anode liquid is returned to the step (2) to be used as a third washing liquid and is returned to the step (1) to be used as a leaching agent.
Example 2
(1) Calcified roasting clinker (containing V) of vanadium slag 2 O 5 13.75%, mnO 7.92%, P0.03%) and a first mother liquor (containing vanadium 4.4g/L, manganese 19.2g/L and phosphorus 0.07g/L, pH =2.4) are stirred and mixed for 5min, wherein the liquid-solid ratio of the first mother liquor to the vanadium slag calcified roasting clinker is 0.9mL/g, then a second mother liquor (containing vanadium 2.4g/L, manganese 11.4g/L and phosphorus 0.03g/L, pH =1.6) is slowly added under stirring for 30min, the pH value of the system in the adding process is controlled to be 2.6-3.5, the liquid-solid ratio of the second mother liquor to the vanadium slag calcified roasting clinker is 0.8mL/g, and then a leaching agent (containing manganese 9.5g/L and phosphorus not more than 0.01g/L, pH =0.8) is slowly added, and the liquid-solid ratio of the leaching agent to the vanadium slag calcified roasting clinker is controlled to be 2.6-3.5The ratio is 0.5mL/g, the leaching is carried out for 40min under the condition of pH value of 2.6 by using sulfuric acid, and residues and leaching liquid (containing 29.8g/L vanadium, 19.3g/L manganese and 0.03g/L phosphorus) are obtained through solid-liquid separation;
(2) Washing the residue obtained in the step (1) by sequentially using a first washing liquid (containing 20.1g/L of manganese and no more than 0.01g/L, pH =6.0), a second washing liquid (containing 0.25g/L of vanadium, 20.9g/L of manganese and 0.02g/L, pH =1.8) of phosphorus and a third washing liquid (containing 9.5g/L of manganese and no more than 0.01g/L, pH =0.8) of phosphorus to respectively obtain a first washing filtrate, a second washing filtrate, a third washing filtrate and vanadium extraction tailings (TV=0.70%); wherein the liquid-solid ratio of the first washing liquid to the vanadium slag calcified roasting clinker is 0.3mL/g, the liquid-solid ratio of the second washing liquid to the vanadium slag calcified roasting clinker is 1.5mL/g, and the liquid-solid ratio of the third washing liquid to the vanadium slag calcified roasting clinker is 0.8mL/g;
(3) Mixing the first washing filtrate with the leaching solution, adding ammonium sulfate for precipitating vanadium to obtain ammonium polyvanadate and vanadium extraction wastewater (containing 0.35g/L of vanadium, 18.5g/L of manganese and 0.02g/L, pH =1.5 of phosphorus), and calcining the ammonium polyvanadate to obtain a vanadium pentoxide product;
(4) Removing impurities from part of vanadium extraction wastewater, neutralizing to obtain impurity-removing solution (vanadium not more than 0.01g/L, manganese not more than 18.8g/L, phosphorus not more than 0.01g/L, pH =6.5), adding electrolytic additive (selenium dioxide) into part of the impurity-removing solution for electrolytic treatment, and controlling current density to 260A/m 2 The concentration of manganese in the cathode region is 8-12g/L, and electrolytic anolyte and manganese metal products are obtained; when the electrolytic additive is added into part of the impurity removing liquid, the volume ratio of the weight of the electrolytic additive to the impurity removing liquid is 0.02g/L;
(5) Returning the second washing filtrate to the step (1) to be used as a first mother liquor, and returning the third washing filtrate to the step (1) to be used as a second mother liquor;
(6) The residual vanadium extraction wastewater is directly returned to the step (2) to be used as a second washing liquid, the residual impurity removal liquid is returned to the step (2) to be used as a first washing liquid, the electrolytic anode liquid is returned to the step (2) to be used as a third washing liquid and is returned to the step (1) to be used as a leaching agent.
Example 3
(1) Calcification roasting of vanadium slagGrog (containing V) 2 O 5 19.08%, mnO9.42%, P0.08%) and a first mother liquor (containing 2.6g/L of vanadium, 19.2g/L of manganese and 0.02g/L, pH =2.0) are mixed for 5min, wherein the liquid-solid ratio of the first mother liquor to the vanadium slag calcified roasting clinker is 1.5mL/g, then a second mother liquor (containing 2.9g/L of vanadium, 9.7g/L of manganese and 0.04g/L, pH =1.0) is slowly added in 30min under stirring, the pH value of the system in the adding process is controlled to be 3.0-3.5, the liquid-solid ratio of the second mother liquor to the vanadium slag calcified roasting clinker is 0.8mL/g, then a leaching agent (containing 8.5g/L of manganese and 0.01g/L, pH =1.0) is slowly added, and the leaching agent and the vanadium slag calcified roasting clinker is slowly added in 0.7mL/g, and the leaching agent and the solid-solid solution is separated from the leaching agent (containing 6.02 g/1.33 g) under the condition that the pH value is 3.2 by using sulfuric acid to obtain 6.02 g/L of vanadium and 6.33 g/L of residues;
(2) Washing the residue obtained in the step (1) by using a first washing liquid (vanadium-containing not more than 0.01g/L, manganese-containing not more than 18.8g/L, phosphorus-containing not more than 0.01g/L, pH =6.5), a second washing liquid (vanadium-containing 0.35g/L, manganese-containing 18.5g/L, phosphorus-containing 0.02g/L, pH =1.5) and a third washing liquid (manganese-containing 8.5g/L, phosphorus-containing not more than 0.01g/L, pH =1.0) in sequence to respectively obtain a first washing filtrate, a second washing filtrate, a third washing filtrate and vanadium-extracting tailings (TV=0.88%); wherein the liquid-solid ratio of the first washing liquid to the vanadium slag calcified roasting clinker is 0.5mL/g, the liquid-solid ratio of the second washing liquid to the vanadium slag calcified roasting clinker is 1.2mL/g, and the liquid-solid ratio of the third washing liquid to the vanadium slag calcified roasting clinker is 1.5mL/g;
(3) Mixing the first washing filtrate with the leaching solution, adding ammonium sulfate for precipitating vanadium to obtain ammonium polyvanadate and vanadium extraction wastewater (containing 0.25g/L of vanadium, 25.4g/L of manganese and 0.02g/L, pH =2.0) and calcining the ammonium polyvanadate to obtain a vanadium pentoxide product;
(4) Removing impurities from part of vanadium extraction wastewater, neutralizing to obtain impurity-removing solution (vanadium not more than 0.01g/L, manganese 24.7g/L, phosphorus not more than 0.01g/L, pH =7.0), adding 0.09g of electrolytic additive (selenium dioxide) into part of impurity-removing solution for electrolytic treatment, and controlling current density to 350A/m 2 The concentration of manganese in the cathode region is 8-12g/L, and electrolytic anolyte and manganese metal products are obtained; when the electrolytic additive is added into part of the impurity removing liquid, the weight of the electrolytic additive and the impurity removing liquidIs 0.02g/L;
(5) Returning the second washing filtrate to the step (1) to be used as a first mother liquor, and returning the third washing filtrate to the step (1) to be used as a second mother liquor;
(6) The residual vanadium extraction wastewater is directly returned to the step (2) to be used as a second washing liquid, the residual impurity removal liquid is returned to the step (2) to be used as a first washing liquid, the electrolytic anode liquid is returned to the step (2) to be used as a third washing liquid and is returned to the step (1) to be used as a leaching agent.
Comparative example 1
1000g (containing V) of vanadium slag calcified roasting clinker 2 O 5 16.06%, mnO 8.64% and P0.05%), adding 2500mL of water, stirring and mixing for 35min, adding sulfuric acid to adjust the pH of the system to 3.0, carrying out acid leaching reaction at normal temperature for 60min, and carrying out solid-liquid separation to obtain leaching liquid (vanadium 31.7g/L, manganese 10.4g/L and phosphorus 0.02 g/L) and residues; the residue was washed three times with water in volumes of 400mL, 1000mL, respectively, to obtain a first washing filtrate, a second washing filtrate, a third washing filtrate, and vanadium extraction tailings (tv=1.23%).
And combining the leaching solution and the first washing filtrate, precipitating vanadium to obtain ammonium polyvanadate and vanadium extraction wastewater (containing 0.15g/L of vanadium, 9.6g/L of manganese and 0.02g/L, pH =1.8 of phosphorus), and calcining to obtain a vanadium pentoxide product. Removing impurities from vanadium extraction wastewater, neutralizing to obtain impurity-removing liquid (vanadium not more than 0.01g/L, manganese not more than 8.9g/L and phosphorus not more than 0.01g/L, pH =6.5), adding selenium dioxide into the impurity-removing liquid, introducing direct current for electrolysis, and controlling current density to 260A/m, wherein the solid-liquid ratio of the selenium dioxide to the impurity-removing liquid is 0.025g/L 2 The manganese concentration in the cathode region is 3-7g/L, and the anolyte (containing 3.9g/L, pH =1.0) and the manganese metal product are obtained.
Comparative example 2:
1000g (containing V) of vanadium slag calcified roasting clinker 2 O 5 16.06%, mnO 8.64% and P0.05%), 2500mL of acid solution (containing 3.9g/L, pH =1.0 of manganese) is added for mixing and pulping, sulfuric acid is added for regulating the pH of the system to 3.0, acid leaching reaction is carried out for 60min at normal temperature, and leaching liquid (containing 27.5g/L of vanadium, 14.2g/L of manganese and 0.08g/L of phosphorus) and residues are obtained through solid-liquid separation; respectively adopt the volume ofThe residue was washed three times with 400mL, 1000mL of solution (manganese containing 3.9g/L, pH =1.0) to give a first washing filtrate, a second washing filtrate, a third washing filtrate and vanadium extraction tailings (tv=1.35%).
And combining the leaching solution and the first washing filtrate, precipitating vanadium to obtain ammonium polyvanadate and vanadium extraction wastewater (containing 0.32g/L of vanadium, 13.9g/L of manganese and 0.04g/L, pH =1.8 of phosphorus), and calcining to obtain a vanadium pentoxide product. Removing impurities from vanadium extraction wastewater, neutralizing to obtain impurity-removing solution (vanadium not more than 0.01g/L, manganese 13.5g/L, phosphorus not more than 0.01g/L, pH =6.0), adding selenium dioxide 0.025g/L into the impurity-removing solution, introducing direct current for electrolysis, and controlling current density to 260A/m 2 The manganese concentration in the cathode region is 3-7g/L, and the anolyte (containing 4.2g/L, pH =1.1) and the manganese metal product are obtained.
Test example 1
The vanadium content in the vanadium extraction tailings obtained in the examples and comparative examples was detected, and the leaching rate of vanadium in the clinker in the leaching process was calculated, and the results are shown in table 1.
The leaching rate calculation formula is as follows: (1-vanadium extraction tailings vanadium content. Tailings weight/(calcified clinker vanadium content. Clinker weight)). 100%
TABLE 1
| Vanadium extraction tailings TV/wt% | Leaching rate/% | |
| Example 1 | 0.95 | 90.6 |
| Example 2 | 0.7 | 91 |
| Example 3 | 0.88 | 92.3 |
| Comparative example 1 | 1.23 | 86.9 |
| Comparative example 2 | 1.35 | 85.2 |
As can be seen from the data in Table 1, the leaching rate in examples 1-3 is significantly higher than that in comparative examples 1-2, which shows that the leaching rate of vanadium in the clinker is higher, reaching 90.6-92.3%.
Test example 2
The main components (wt%) of the vanadium pentoxide products obtained in examples and comparative examples were examined, and the results are shown in table 2; the main components (wt%) of the manganese metal products obtained in examples and comparative examples were examined, and the results are shown in table 3.
TABLE 2
TABLE 3 Table 3
From the data in tables 2 and 3, it can be seen that the vanadium pentoxide product and the manganese metal product obtained in the examples meet the requirements of YBT5304-2017 and YB/T051-2015, respectively, which shows that the vanadium pentoxide product and the manganese metal product meeting the quality requirements can be obtained by adopting the method of the invention. However, in comparative example 2, the solution with lower acidity (ph=1.0) was directly used for clinker pulping, vanadium was rapidly dissolved into the solution, and when the vanadium concentration of the solution was higher (e.g. vanadium concentration was higher than 5 g/L) and the pH was still lower (e.g. lower than 1.5), precipitation of vanadium occurred to cause the increase of residue TV (see table 1); meanwhile, phosphorus impurities also quickly enter the solution under the low pH condition, vanadium-phosphorus heteropolyacid is formed by the phosphorus impurities and dissolved vanadium, the phosphorus content in the obtained leaching solution is up to 0.08g/L, and the P content of a vanadium pentoxide product prepared by adopting the leaching solution is up to 0.035% and exceeds the standard requirement.
Test example 3
The current efficiency of the process of preparing the metal manganese product by electrolyzing the impurity removing liquid in the examples and the comparative examples is calculated, and the results are shown in Table 4.
TABLE 4 Table 4
| Current efficiency/% | |
| Example 1 | 75.2 |
| Example 2 | 73.6 |
| Example 3 | 77.1 |
| Comparative example 1 | 46.4 |
| Comparative example 2 | 55.7 |
From the data in Table 4, the current efficiency of the electrolytic recovery of metal manganese in the examples is significantly higher than that of the comparative examples, which shows that by adopting the method of the invention, the manganese-containing solution is recycled back to the clinker leaching mode to increase the manganese concentration of the system, thereby effectively improving the electrolytic efficiency, reducing the electrolytic energy consumption, reducing the loss and reducing the wastewater discharge.
The preferred embodiments of the present invention have been described in detail above, but the present invention is not limited thereto. Within the scope of the technical idea of the invention, a number of simple variants of the technical solution of the invention are possible, including combinations of the individual technical features in any other suitable way, which simple variants and combinations should likewise be regarded as being disclosed by the invention, all falling within the scope of protection of the invention.
Claims (11)
1. The method for extracting vanadium, manganese and recycling waste water from the vanadium slag calcified roasting clinker is characterized by comprising the following steps of:
(1) Mixing the vanadium slag calcified roasting clinker with the first mother liquor under stirring, adding the second mother liquor under stirring, adding a leaching agent, and leaching under the condition that the pH value is 2.6-3.5 to obtain residues and leaching liquid;
(2) Washing residues obtained in the step (1) sequentially by using a first washing liquid, a second washing liquid and a third washing liquid to obtain a first washing filtrate, a second washing filtrate, a third washing filtrate and vanadium extraction tailings respectively;
(3) Mixing the first washing filtrate with the leaching solution, adding ammonium salt for precipitating vanadium to obtain ammonium polyvanadate and vanadium extraction wastewater, and calcining the ammonium polyvanadate to obtain a vanadium pentoxide product;
(4) Removing impurities from part of vanadium extraction wastewater, neutralizing to obtain an impurity-removing solution, and then adding an electrolysis additive into part of the impurity-removing solution for electrolysis treatment to obtain an electrolysis anolyte and a metal manganese product;
(5) Returning the second washing filtrate to the step (1) to be used as a first mother liquor, and returning the third washing filtrate to the step (1) to be used as a second mother liquor;
(6) Returning the residual vanadium extraction wastewater to the step (2), returning the residual impurity removal liquid to the step (2), and returning the electrolytic anolyte to the step (2) and/or the step (1);
in the step (1), the vanadium slag calcified roasting clinker contains 12-25 wt% of V 2 O 5 5 to 10 weight percent of MnO and less than or equal to 0.1 weight percent of P,
the first mother solution is manganese-containing solution with pH=1.5-8 and P less than or equal to 0.08 g/L;
the second mother solution is manganese-containing solution with P less than or equal to 0.08 g/L;
in the step (4), controlling the manganese concentration of the cathode region to be 8-15g/L in the electrolytic treatment process, and controlling the current density to be 260-350A/m in the electrolytic treatment process 2 。
2. The method for extracting vanadium, manganese and recycling waste water from vanadium slag calcified roasting clinker according to claim 1, wherein in step (1), the liquid-solid ratio of the first mother liquor to the vanadium slag calcified roasting clinker is 0.6-2mL/g.
3. The method for extracting vanadium, manganese and recycling waste water from vanadium slag calcified roasting clinker according to claim 2, wherein the liquid-solid ratio of the second mother liquor to the vanadium slag calcified roasting clinker is 0.8-1.5mL/g.
4. The method for extracting vanadium, manganese and recycling waste water from vanadium slag calcified roasting clinker according to claim 2, wherein the liquid-solid ratio of the leaching agent to the vanadium slag calcified roasting clinker is 0.5-1.5mL/g.
5. The method for extracting vanadium, manganese and recycling waste water from vanadium slag calcified roasting clinker according to claim 2, wherein the ratio of the total volume of the first mother liquor, the second mother liquor and the leaching agent to the weight of the vanadium slag calcified roasting clinker is 2-4mL/g.
6. The method for extracting vanadium, manganese and recycling waste water from vanadium slag calcified roasting clinker according to claim 1, wherein in step (1), the leaching time is 40-60min.
7. The method for extracting vanadium, manganese and recycling waste water from vanadium slag calcified roasting clinker according to claim 1, wherein in step (1), the leaching solution contains 15-40g/L manganese.
8. The method for extracting vanadium, manganese and recycling waste water from vanadium slag calcified roasting clinker according to claim 1, wherein in the step (2), the liquid-solid ratio of the first washing liquid to the vanadium slag calcified roasting clinker is 0.2-0.5mL/g.
9. The method for extracting vanadium, manganese and recycling waste water from vanadium slag calcified roasting clinker according to claim 1, wherein in the step (2), the liquid-solid ratio of the second washing liquid to the vanadium slag calcified roasting clinker is 0.6-2mL/g.
10. The method for extracting vanadium, manganese and wastewater recycling of the vanadium slag calcified roasting clinker according to claim 1 or 9, wherein in the step (2), the liquid-solid ratio of the third washing liquid to the vanadium slag calcified roasting clinker is 0.8-1.5mL/g.
11. The method for extracting vanadium, manganese and recycling waste water from the vanadium slag calcified roasting clinker according to claim 1, wherein in the step (6), when the pH=1.2-4 and the P content of the residual vanadium extraction waste water is less than or equal to 0.08g/L, the residual vanadium extraction waste water is directly returned to the step (2) for use; when the vanadium extraction wastewater does not meet the requirements, the wastewater needs to be adjusted to be within the required range, and then the wastewater returns to the step (2) for use.
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