CN111705228A - Method for mixed precipitation of tetra-and pentavalent vanadium - Google Patents
Method for mixed precipitation of tetra-and pentavalent vanadium Download PDFInfo
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- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
- C22B3/44—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes
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- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
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Abstract
The invention belongs to the technical field of vanadium hydrometallurgy, and particularly discloses a method for mixed precipitation of tetra-valent vanadium and pentavalent vanadium, which comprises the following steps: heating the vanadium-containing leaching solution to 70-100 ℃, adding a reducing agent to reduce part of vanadium, then cooling to 20-50 ℃, adjusting the pH of the solution to 4.5-5.5 by using ammonium-containing carbonate, performing solid-liquid separation to obtain vanadium-precipitated supernatant and vanadium-containing precipitate, pulping, washing and calcining the vanadium-containing precipitate by using ammonium bicarbonate to obtain vanadium pentoxide. The method can reduce the consumption of the vanadium reduction reagent, the supernatant can be directly recycled, and the generation of solid wastes such as vanadium-chromium reduction filter cakes, ammonium-containing sodium sulfate and the like in the existing vanadium oxide production process is avoided.
Description
Technical Field
The invention belongs to the technical field of vanadium hydrometallurgy, and particularly relates to a method for mixed precipitation of tetra-valent vanadium and pentavalent vanadium.
Background
Vanadium slag is a main raw material for preparing vanadium oxide, the traditional production process is sodium roasting-water extraction of vanadium, in the vanadium extraction process, sodium oxide vanadium extraction tailings containing about 6 percent can be generated, the vanadium extraction tailings are difficult to be reused, vanadium-chromium reduction filter cakes and a large amount of solid waste sodium sulfate are difficult to treat, and the environmental protection hidden danger is large; a large amount of sodium carbonate is consumed in the vanadium extraction process, and the process cost is high. In order to reduce the production cost of vanadium oxide and eliminate the hidden trouble of environmental protection, the process idea of preparing vanadium oxide by calcifying roasting-carbonating leaching of vanadium slag is provided.
In the literature, "research on roasting-leaching reaction process mechanism of high-calcium low-grade vanadium slag" the concentration of sodium carbonate used as a leaching agent in the research on the roasting and sodium carbonate leaching process of vanadium slag is 160g/L, and the solid-to-solid ratio of a leaching solution is 10: 1 (mL/g). The leaching solution obtained by the method has low vanadium concentration and high Na/V ratio, so that the vanadium precipitation rate is low when ammonium metavanadate is subsequently precipitated.
CN102560086A discloses a vanadium extraction method for leaching vanadium slag clinker with ammonium carbonate, which comprises the following steps of adopting 200-800 g/L of ammonium carbonate solution and vanadium slag clinker according to a liquid-solid ratio of 5: 1-30: 1 leaching at 60-98 ℃. The method has the advantages of high leaching agent consumption and high production cost; in addition, because the solubility of ammonium metavanadate is low, in order to avoid vanadium from precipitating into residues in the form of ammonium metavanadate, the method needs to control a larger liquid-solid ratio when ammonium carbonate is used for leaching, so that the concentration of vanadium in the obtained leaching solution is lower; on the other hand, the solubility of ammonium metavanadate is reduced along with the reduction of temperature, and the ammonium metavanadate is easily separated out when the temperature of the leaching solution is reduced by the method, so that the solution system is unstable.
Disclosure of Invention
The invention aims to solve the problems of large solid waste amount, high sodium-vanadium ratio of carbonated leachate and large consumption of ammonium salt in the existing vanadium oxide production technology.
In order to solve the technical problem, the invention provides a method for mixing and precipitating tetravalence vanadium and pentavalence vanadium, which comprises the following steps: heating the vanadium-containing leaching solution to 70-100 ℃, adding a reducing agent to reduce part of vanadium, then cooling to 20-50 ℃, adjusting the pH of the solution to 4.5-5.5 by using ammonium-containing carbonate, performing solid-liquid separation to obtain vanadium-precipitated supernatant and vanadium-containing precipitate, pulping, washing and calcining the vanadium-containing precipitate by using ammonium bicarbonate to obtain vanadium pentoxide.
In the method for mixing and precipitating the tetravalence vanadium and the pentavalence vanadium, the reducing agent is oxalic acid + ammonium oxalate, tartaric acid or ascorbic acid; the reduction time is 30 +/-5 min.
In the method for mixing and precipitating the tetravalence vanadium and the pentavalence vanadium, when the reducing agent is oxalic acid and ammonium oxalate, at least one of the following conditions is met:
the temperature of the vanadium-containing leaching solution participating in the reduction reaction is 70 +/-3 ℃;
adding oxalic acid according to the pH value of the solution controlled to be 2.0-3.5; the dosage of the ammonium oxalate is controlled according to the molar ratio of total oxalate ions added into the solution to the total vanadium in the leaching solution being 0.2-0.3.
In the method for mixing and precipitating the tetravalence vanadium and the pentavalence vanadium, when the reducing agent is tartaric acid, at least one of the following conditions is met:
the temperature of the vanadium-containing leaching solution participating in the reduction reaction is 80-100 ℃;
the dosage of the tartaric acid is controlled according to the molar ratio of the tartaric acid to the total vanadium in the leaching solution of 0.08-0.12.
In the method for mixing and precipitating the tetravalence vanadium and the pentavalence vanadium, when the reducing agent is ascorbic acid, at least one of the following conditions is met:
the temperature of the vanadium-containing leaching solution participating in the reduction reaction is 80-100 ℃;
the dosage of the ascorbic acid is controlled according to the molar ratio of the ascorbic acid to the total vanadium in the leaching solution of 0.03-0.06.
In the method for mixing and precipitating the tetravalence vanadium and the pentavalence vanadium, the ammonium-containing carbonate is ammonium carbonate and/or ammonium bicarbonate.
In the method for mixed precipitation of the tetravalence vanadium and the pentavalence vanadium, the leaching solution containing vanadium is subjected to impurity removal by using sodium aluminate before use, and the use amount of the sodium aluminate is controlled according to the molar ratio of Al/Si of 0.6-2.0.
According to the method for mixing and precipitating the tetravalence vanadium and the pentavalence vanadium, the vanadium-containing leaching solution is prepared by the following method: mixing vanadium slag and calcium salt uniformly, and roasting at 800-950 ℃ for 40-200 min to obtain roasted clinker; and (3) stirring and leaching the roasted clinker for 30-150 min by using a sodium-containing carbonate solution at 80-100 ℃, and performing solid-liquid separation to obtain a vanadium-containing leaching solution, wherein the amount of the calcium salt is 0-8% of the weight of the vanadium slag by taking CaO as an amount.
Wherein, the method for mixing and precipitating the tetravalence vanadium and the pentavalence vanadium meets at least one of the following conditions:
the calcium salt is more than one of calcium carbonate, calcium hydroxide or calcium oxide;
the sodium-containing carbonate is sodium carbonate and/or sodium bicarbonate;
the sodium carbonate is used in an amount of CO3 2-The molar weight of vanadium in the roasting clinker is 1.5-3.5 times;
when leaching, the ratio of the solution to the roasted clinker is 1.8-3.0 mL: 1g of the total weight of the composition.
In the method for mixing and precipitating the tetravalence vanadium and the pentavalence vanadium, the supernatant fluid of the precipitated vanadium is directly recycled for leaching the roasted clinker.
Compared with the prior art, the invention has the beneficial effects that:
according to the method, a reducing agent is added into the solution after impurity removal to reduce part of pentavalent vanadium and vanadium is separated in a mode of precipitating tetravalent vanadium and sodium ammonium vanadate, so that the consumption of a vanadium reduction reagent is reduced. The method of the invention can not generate vanadium-chromium reduction filter cakes and solid waste sodium sulfate, and the obtained vanadium precipitation supernatant can be returned to the leaching process to be used as a leaching agent for recycling, thereby realizing low-cost clean production of vanadium oxide by vanadium slag, reducing the consumption of the leaching agent, and solving the problems of high process cost and difficult utilization of the solid waste vanadium-chromium reduction filter cakes and sodium sulfate in the traditional vanadium slag sodium salt roasting-water leaching of vanadium.
Detailed Description
Specifically, the method for mixed precipitation of tetravalence vanadium and pentavalence vanadium comprises the following steps: heating the vanadium-containing leaching solution to 70-100 ℃, adding a reducing agent to reduce part of vanadium, then cooling to 20-50 ℃, adjusting the pH of the solution to 4.5-5.5 by using ammonium-containing carbonate, performing solid-liquid separation to obtain vanadium-precipitated supernatant and vanadium-containing precipitate, pulping, washing and calcining the vanadium-containing precipitate by using ammonium bicarbonate to obtain vanadium pentoxide.
In the method, the reducing agent is adopted to reduce the vanadium part so as to reduce the dosage of the reducing agent and reduce the cost of the agent; the reason that the ammonium carbonate is selected to adjust the pH value to be 4.5-5.5 is that the ammonium carbonate does not cause the concentration change of sodium ions in a solution system while adjusting the pH value, and the ammonium radical can also precipitate pentavalent vanadium to form ammonium vanadateSodium precipitation, and realizes the synchronous precipitation of tetravalent vanadium and pentavalent vanadium. Specifically, when the reducing agent is oxalic acid and ammonium oxalate, the oxalic acid can adjust the pH of the solution to acidity; meanwhile, oxalic acid and oxalate in ammonium oxalate can reduce part of pentavalent vanadium into tetravalent vanadium, and ammonium can be used for precipitating pentavalent vanadium after the pH value of the solution is adjusted to be 4.5-5.5; control of total C in solution2O4 2-The reason why the molar ratio of V/V is 0.2 to 0.3 is to reduce the amount of the reducing agent used. When the reducing agent is tartaric acid or ascorbic acid, the pH of the solution can be adjusted to be acidic, and meanwhile, pentavalent vanadium can be reduced into tetravalent vanadium by the tartaric acid and the ascorbic acid; controlling C in solution4H6O60.08 to 0.12 (molar ratio) or C6H8O6The reason why the/V is 0.03 to 0.06 (molar ratio) is to reduce part of pentavalent vanadium to tetravalent vanadium, thereby reducing the cost of the reducing agent.
According to the method, before the vanadium-containing leachate is used, sodium aluminate is used for removing silicon and phosphorus in the leachate at the same time, and the use amount of the sodium aluminate is controlled according to the molar ratio of Al/Si of 0.6-2.0.
In the method, the vanadium-containing leaching solution is prepared by the following steps: uniformly mixing vanadium slag and calcium salt accounting for 0-8% of the weight of the vanadium slag, and roasting at 800-950 ℃ for 40-200 min to obtain roasted clinker; and stirring and leaching the roasted clinker with a sodium-containing carbonate solution at 80-100 ℃ for 30-150 min, and carrying out solid-liquid separation to obtain a vanadium-containing leaching solution.
The vanadium slag is common vanadium slag or high-calcium high-phosphorus vanadium slag obtained by oxidizing and blowing vanadium-containing molten iron. In order to fully expose the ferrovanadium spinel in the vanadium slag, facilitate the oxidation of the ferrovanadium spinel and the full contact reaction of the vanadium slag and calcium salt, crushing the vanadium slag, and selecting the vanadium slag with the granularity of less than 0.096 mm. Because the solubility of calcium metavanadate in water is larger than that of calcium pyrovanadate and calcium orthovanadate, the calcium metavanadate is beneficial to leaching, and in order to control the vanadium in the calcified roasting clinker to mainly exist in the form of calcium metavanadate, the amount of added calcium salt is 0-8 percent of the mass of the vanadium slag calculated by CaO.
The roasted clinker is leached by sodium-containing carbonate, and the carbonation leaching is adopted because the solubility of calcium carbonate is higher than that of calcium metavanadate, calcium pyrovanadate and normal calciumThe calcium vanadate is small, calcium ions and carbonate are combined into slag, and vanadium enters solution. The inventor controls the dosage of the sodium carbonate to be CO3 2-the/V is 1.5-3.5 (mol ratio) to provide enough carbonate for leaching, and the leaching rate of vanadium is improved. In order to further obtain high-concentration vanadium-containing leachate, the solid-to-solid ratio of the leachate is controlled to be 1.8-2.5: 1 (mL: g). In order to facilitate the dissolution of calcium metavanadate, the leaching temperature is controlled to be 80-100 ℃.
In the method, the vanadium-containing precipitate is pulped and washed by using the ammonium bicarbonate solution, so that the sodium ammonium vanadate is converted into the ammonium metavanadate, and the sodium content in the vanadium product is reduced.
In the method, the supernatant fluid of the precipitated vanadium can be directly recycled, and the generation of solid wastes such as vanadium-chromium reduction filter cakes, ammonium-containing sodium sulfate and the like in the existing production process of vanadium oxide is avoided.
The following examples are provided to further illustrate the embodiments of the present invention and are not intended to limit the scope of the present invention.
Example 1
Taking vanadium slag (containing V) with the granularity less than 0.096mm2O517.2%, CaO1.84%, P0.04%) was calcined in a muffle furnace at 950 ℃ for 40min with air. Crushing the roasted clinker, adding the crushed roasted clinker into 180mL of water, adding 50g of sodium carbonate, stirring and leaching for 120min at the slurry temperature of 95 ℃, wherein the solid-to-solid ratio of a leaching solution is 1.8: 1 (mL: g), solid-liquid separation to obtain leachate and residue, wherein the residue TV0.92wt% and Na0.39%, and the vanadium leaching rate is 90.3%.
Adding 0.8g of sodium aluminate into the leaching solution, stirring for 20min, and filtering to obtain a solution after impurity removal.
Adjusting the pH of the solution after impurity removal to 2.0 by using oxalic acid according to the total C in the solution2O4 2-Adding ammonium oxalate with the molar ratio of 0.2 to V, and stirring and reacting for 30min at 70 ℃; cooling the solution to 30 ℃, adjusting the pH value to 5.0 by using ammonium bicarbonate, and carrying out solid-liquid separation to obtain a vanadium-containing precipitate and a vanadium precipitation supernatant.
Pulping the vanadium-containing precipitate with ammonium bicarbonate solution, washing, and oxidizing at 500 deg.CCalcining for 3h to obtain 15.36g of vanadium pentoxide product, wherein V2O5The content is 98.6%; the supernatant of the precipitated vanadium is used as a leaching agent to be directly used for the next leaching.
Example 2
Taking vanadium slag (containing V) with the granularity less than 0.096mm2O517.2 percent, CaO1.84 percent and P0.04 percent) and 3g of calcium oxide are evenly mixed and roasted for 80min by a muffle furnace under the conditions of roasting temperature of 900 ℃ and air introduction. Crushing the roasted clinker, adding the crushed roasted clinker into 220mL of water, simultaneously adding 55g of sodium bicarbonate, stirring and leaching for 150min at the slurry temperature of 85 ℃, wherein the solid-to-solid ratio of a leaching solution is 2.1: 1 (mL: g), solid-liquid separation to obtain vanadium-containing leachate and residue, wherein the residue TV0.90wt% and Na0.34%, and the vanadium leaching rate is 90.4%.
Adding 0.7g of sodium aluminate into the leaching solution, stirring for 20min, and filtering to obtain a solution after impurity removal.
Adjusting the pH of the solution after impurity removal to 2.5 by using oxalic acid according to the total C in the solution2O4 2-Adding ammonium oxalate with the molar ratio of 0.25 to V, and stirring and reacting for 30min at 70 ℃; cooling the solution to 20 ℃, adjusting the pH value to 4.8 by using ammonium bicarbonate, and carrying out solid-liquid separation to obtain a vanadium-containing precipitate and a vanadium precipitation supernatant.
Pulping and washing the vanadium-containing precipitate by ammonium bicarbonate solution, and oxidizing and calcining the vanadium-containing precipitate for 3 hours at 500 ℃ to obtain 15.40g of vanadium pentoxide product, wherein V2O5The content is 98.4%; the supernatant of the precipitated vanadium is used as a leaching agent to be directly used for the next leaching.
Example 3
Taking vanadium slag (containing V) with the granularity less than 0.096mm2O517.2 percent, CaO1.84 percent and P0.04 percent) and 8g of calcium oxide are evenly mixed and roasted for 180min by a muffle furnace under the conditions of roasting temperature of 850 ℃ and air introduction; crushing the roasted clinker, adding the crushed roasted clinker into 250mL of water, adding 64g of sodium carbonate, stirring and leaching for 100min at the slurry temperature of 90 ℃, wherein the solid-to-solid ratio of a leaching solution is 2.3: 1 (mL: g), solid-liquid separation to obtain leachate and residue, wherein the residue TV0.92wt% and Na0.37%, and the vanadium leaching rate is 90.3%.
Adding 0.8g of sodium aluminate into the leaching solution, stirring for 20min, and filtering to obtain a solution after impurity removal.
Adjusting the pH value of the solution after impurity removal to be 3.0 by using oxalic acid according to the total C in the solution2O4 2-Adding ammonium oxalate with the molar ratio of 0.3 to V, and stirring and reacting for 30min at 70 ℃; cooling the solution to 40 ℃, adjusting the pH value to 5.5 by using ammonium carbonate, and carrying out solid-liquid separation to obtain vanadium-containing precipitate and vanadium precipitation supernatant.
Pulping and washing the vanadium-containing precipitate by ammonium bicarbonate solution, and oxidizing and calcining the vanadium-containing precipitate for 3 hours at 500 ℃ to obtain 15.38g of vanadium pentoxide product, wherein V2O5The content is 98.5%; the supernatant of the precipitated vanadium is used as a leaching agent to be directly used for the next leaching.
Example 4
Taking vanadium slag (containing V) with the granularity less than 0.096mm2O517.2%, CaO1.84%, P0.04%) was calcined in a muffle furnace at 950 ℃ for 40min with air. Crushing the roasted clinker, adding the crushed roasted clinker into 180mL of water, adding 50g of sodium carbonate, stirring and leaching for 120min at the slurry temperature of 95 ℃, wherein the solid-to-solid ratio of a leaching solution is 1.8: 1 (mL: g), solid-liquid separation to obtain leachate and residue, wherein the residue TV0.92wt% and Na0.39%, and the vanadium leaching rate is 90.3%.
Adding 0.8g of sodium aluminate into the leaching solution, stirring for 20min, and filtering to obtain a solution after impurity removal.
According to C4H6O6Adding tartaric acid into the solution after impurity removal, and stirring and reacting for 30min at 80 ℃; cooling the solution to 30 ℃, adjusting the pH value to 5.0 by using ammonium bicarbonate, and carrying out solid-liquid separation to obtain a vanadium-containing precipitate and a vanadium precipitation supernatant.
Pulping and washing the vanadium-containing precipitate by ammonium bicarbonate solution, and oxidizing and calcining the vanadium-containing precipitate for 3 hours at 500 ℃ to obtain 15.36g of vanadium pentoxide product, wherein V2O5The content is 98.6%; the supernatant of the precipitated vanadium is used as a leaching agent to be directly used for the next leaching.
Example 5
Taking vanadium slag (containing V) with the granularity less than 0.096mm2O517.2 percent, CaO1.84 percent and P0.04 percent) and 3g of calcium oxide are evenly mixed and roasted for 80min by a muffle furnace under the conditions of roasting temperature of 900 ℃ and air introduction. Adding the roasted clinker after crushingAdding 55g of sodium bicarbonate into 220mL of water, and stirring and leaching for 150min at the slurry temperature of 85 ℃, wherein the solid-to-solid ratio of a leaching solution is 2.1: 1 (mL: g), solid-liquid separation to obtain vanadium-containing leachate and residue, wherein the residue TV0.90wt% and Na0.34%, and the vanadium leaching rate is 90.4%.
Adding 0.7g of sodium aluminate into the leaching solution, stirring for 20min, and filtering to obtain a solution after impurity removal.
According to C4H6O60.10 (molar ratio), adding tartaric acid into the solution after impurity removal, and stirring and reacting for 30min at 90 ℃; cooling the solution to 20 ℃, adjusting the pH value to 4.8 by using ammonium bicarbonate, and carrying out solid-liquid separation to obtain a vanadium-containing precipitate and a vanadium precipitation supernatant.
Pulping and washing the vanadium-containing precipitate by ammonium bicarbonate solution, and oxidizing and calcining the vanadium-containing precipitate for 3 hours at 500 ℃ to obtain 15.41g of vanadium pentoxide product, wherein V2O5The content is 98.4%; the supernatant of the precipitated vanadium is used as a leaching agent to be directly used for the next leaching.
Example 6
Taking vanadium slag (containing V) with the granularity less than 0.096mm2O517.2 percent, CaO1.84 percent and P0.04 percent) and 8g of calcium oxide are evenly mixed and roasted for 180min by a muffle furnace under the conditions of roasting temperature of 850 ℃ and air introduction; crushing the roasted clinker, adding the crushed roasted clinker into 250mL of water, adding 64g of sodium carbonate, stirring and leaching for 100min at the slurry temperature of 90 ℃, wherein the solid-to-solid ratio of a leaching solution is 2.3: 1 (mL: g), solid-liquid separation to obtain leachate and residue, wherein the residue TV0.92wt% and Na0.37%, and the vanadium leaching rate is 90.3%.
Adding 0.8g of sodium aluminate into the leaching solution, stirring for 20min, and filtering to obtain a solution after impurity removal.
According to C4H6O6Adding tartaric acid into the solution after impurity removal, and stirring and reacting for 30min at 100 ℃; cooling the solution to 40 ℃, adjusting the pH value to 5.5 by using ammonium carbonate, and carrying out solid-liquid separation to obtain vanadium-containing precipitate and vanadium precipitation supernatant.
Pulping and washing the vanadium-containing precipitate by ammonium bicarbonate solution, and oxidizing and calcining the vanadium-containing precipitate for 3 hours at 500 ℃ to obtain 15.38g of vanadium pentoxide product, wherein V2O5The content is 98.5%; the supernatant of precipitated vanadium is used as leaching agentDirectly used for the next round of leaching.
Example 7
Taking vanadium slag (containing V) with the granularity less than 0.096mm2O517.2%, CaO1.84%, P0.04%) was calcined in a muffle furnace at 950 ℃ for 40min with air. Crushing the roasted clinker, adding the crushed roasted clinker into 180mL of water, adding 50g of sodium carbonate, stirring and leaching for 120min at the slurry temperature of 95 ℃, wherein the solid-to-solid ratio of a leaching solution is 1.8: 1 (mL: g), solid-liquid separation to obtain leachate and residue, wherein the residue TV0.92wt% and Na0.39%, and the vanadium leaching rate is 90.3%.
Adding 0.8g of sodium aluminate into the leaching solution, stirring for 20min, and filtering to obtain a solution after impurity removal.
According to C6H8O6Adding ascorbic acid into the solution after impurity removal, and stirring and reacting at 80 ℃ for 30 min; cooling the solution to 30 ℃, adjusting the pH value to 5.0 by using ammonium bicarbonate, and carrying out solid-liquid separation to obtain a vanadium-containing precipitate and a vanadium precipitation supernatant.
Pulping and washing the vanadium-containing precipitate by ammonium bicarbonate solution, and oxidizing and calcining the vanadium-containing precipitate for 3 hours at 500 ℃ to obtain 15.36g of vanadium pentoxide product, wherein V2O5The content is 98.6%; the supernatant of the precipitated vanadium is used as a leaching agent to be directly used for the next leaching.
Example 8
Taking vanadium slag (containing V) with the granularity less than 0.096mm2O517.2 percent, CaO1.84 percent and P0.04 percent) and 3g of calcium oxide are evenly mixed and roasted for 80min by a muffle furnace under the conditions of roasting temperature of 900 ℃ and air introduction. Crushing the roasted clinker, adding the crushed roasted clinker into 220mL of water, simultaneously adding 55g of sodium bicarbonate, stirring and leaching for 150min at the slurry temperature of 85 ℃, wherein the solid-to-solid ratio of a leaching solution is 2.1: 1 (mL: g), solid-liquid separation to obtain vanadium-containing leachate and residue, wherein the residue TV0.90wt% and Na0.34%, and the vanadium leaching rate is 90.4%.
Adding 0.7g of sodium aluminate into the leaching solution, stirring for 20min, and filtering to obtain a solution after impurity removal.
According to C6H8O6Adding ascorbic acid into the solution after impurity removal, and stirring and reacting for 30min at 90 ℃; cooling the solution to 20 deg.CAdjusting pH to 4.8 with ammonium bicarbonate, and performing solid-liquid separation to obtain vanadium-containing precipitate and vanadium precipitation supernatant.
Pulping and washing the vanadium-containing precipitate by ammonium bicarbonate solution, and oxidizing and calcining the vanadium-containing precipitate for 3 hours at 500 ℃ to obtain 15.41g of vanadium pentoxide product, wherein V2O5The content is 98.4%; the supernatant of the precipitated vanadium is used as a leaching agent to be directly used for the next leaching.
Example 9
Taking vanadium slag (containing V) with the granularity less than 0.096mm2O517.2 percent, CaO1.84 percent and P0.04 percent) and 8g of calcium oxide are evenly mixed and roasted for 180min by a muffle furnace under the conditions of roasting temperature of 850 ℃ and air introduction; crushing the roasted clinker, adding the crushed roasted clinker into 250mL of water, adding 64g of sodium carbonate, stirring and leaching for 100min at the slurry temperature of 90 ℃, wherein the solid-to-solid ratio of a leaching solution is 2.3: 1 (mL: g), solid-liquid separation to obtain leachate and residue, wherein the residue TV0.92wt% and Na0.37%, and the vanadium leaching rate is 90.3%.
Adding 0.8g of sodium aluminate into the leaching solution, stirring for 20min, and filtering to obtain a solution after impurity removal.
According to C6H8O6Adding ascorbic acid into the solution after impurity removal, and stirring and reacting for 30min at 100 ℃; cooling the solution to 40 ℃, adjusting the pH value to 5.5 by using ammonium carbonate, and carrying out solid-liquid separation to obtain vanadium-containing precipitate and vanadium precipitation supernatant.
Pulping and washing the vanadium-containing precipitate by ammonium bicarbonate solution, and oxidizing and calcining the vanadium-containing precipitate for 3 hours at 500 ℃ to obtain 15.38g of vanadium pentoxide product, wherein V2O5The content is 98.5%; the supernatant of the precipitated vanadium is used as a leaching agent to be directly used for the next leaching.
Claims (10)
1. The method for mixing and precipitating the tetravalence vanadium and the pentavalence vanadium is characterized by comprising the following steps of: heating the vanadium-containing leaching solution to 70-100 ℃, adding a reducing agent to reduce part of vanadium, then cooling to 20-50 ℃, adjusting the pH of the solution to 4.5-5.5 by using ammonium-containing carbonate, performing solid-liquid separation to obtain vanadium-precipitated supernatant and vanadium-containing precipitate, pulping, washing and calcining the vanadium-containing precipitate by using ammonium bicarbonate to obtain vanadium pentoxide.
2. The method for mixed precipitation of tetra-and pentavalent vanadium according to claim 1, wherein the reducing agent is oxalic acid + ammonium oxalate, tartaric acid or ascorbic acid; the reduction time is 30 +/-5 min.
3. The method for mixed precipitation of tetra-and pentavalent vanadium according to claim 2, wherein when the reducing agent is oxalic acid + ammonium oxalate, at least one of the following conditions is satisfied:
the temperature of the vanadium-containing leaching solution participating in the reduction reaction is 70 +/-3 ℃;
adding oxalic acid according to the pH value of the solution controlled to be 2.0-3.5; the dosage of the ammonium oxalate is controlled according to the molar ratio of total oxalate ions added into the solution to the total vanadium in the leaching solution being 0.2-0.3.
4. The method for mixed precipitation of tetra-and pentavalent vanadium according to claim 2, wherein the reducing agent is tartaric acid, and at least one of the following is satisfied:
the temperature of the vanadium-containing leaching solution participating in the reduction reaction is 80-100 ℃;
the dosage of the tartaric acid is controlled according to the molar ratio of the tartaric acid to the total vanadium in the leaching solution of 0.08-0.12.
5. The method for mixed precipitation of tetra-and pentavalent vanadium according to claim 2, wherein the reducing agent is ascorbic acid, and at least one of the following is satisfied:
the temperature of the vanadium-containing leaching solution participating in the reduction reaction is 80-100 ℃;
the dosage of the ascorbic acid is controlled according to the molar ratio of the ascorbic acid to the total vanadium in the leaching solution of 0.03-0.06.
6. The method for mixed precipitation of tetra-and pentavalent vanadium according to any one of claims 1 to 5, wherein the ammonium-containing carbonate is ammonium carbonate and/or ammonium bicarbonate.
7. The method for mixing and precipitating the tetravalence and pentavalence vanadium as claimed in claim 6, wherein the leachate containing vanadium is purified by sodium aluminate before use, and the amount of the sodium aluminate is controlled according to the molar ratio of Al/Si of 0.6-2.0.
8. The method for mixed precipitation of tetravalent and pentavalent vanadium according to claim 7, wherein the vanadium-containing leachate is prepared by the following method: mixing vanadium slag and calcium salt uniformly, and roasting at 800-950 ℃ for 40-200 min to obtain roasted clinker; and (3) stirring and leaching the roasted clinker for 30-150 min by using a sodium-containing carbonate solution at 80-100 ℃, and performing solid-liquid separation to obtain a vanadium-containing leaching solution, wherein the amount of the calcium salt is 0-8% of the weight of the vanadium slag by taking CaO as an amount.
9. The method for mixed precipitation of tetra-and pentavalent vanadium according to claim 8, characterized in that at least one of the following is satisfied:
the calcium salt is more than one of calcium carbonate, calcium hydroxide or calcium oxide;
the sodium-containing carbonate is sodium carbonate and/or sodium bicarbonate;
the sodium carbonate is used in an amount of CO3 2-The molar weight of vanadium in the roasting clinker is 1.5-3.5 times;
when leaching, the ratio of the solution to the roasted clinker is 1.8-3.0 mL: 1g of the total weight of the composition.
10. The method for mixed precipitation of tetra-and pentavalent vanadium according to claim 8 or 9, characterized in that the vanadium precipitation supernatant is directly recycled for leaching the roasted clinker.
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