CN114350981B - Method for recovering vanadium from calcified vanadium extraction tailings - Google Patents
Method for recovering vanadium from calcified vanadium extraction tailings Download PDFInfo
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Abstract
The invention relates to the technical field of extraction metallurgy, and discloses a method for recovering vanadium from calcified vanadium extraction tailings. The method comprises the following steps: (1) Pulping calcified vanadium extraction tailings and calcium sulfite in water, adding sulfuric acid for leaching, adjusting the pH value of the system to 2-3 by lime and/or limestone after the leaching is finished, and then carrying out solid-liquid separation to obtain residues and leaching liquid; (2) Adjusting the pH value of the leaching solution to 5.5-7 by using magnesium carbonate and/or manganese carbonate to precipitate vanadium, and then filtering to obtain vanadium precipitate and wastewater; (3) Adding sodium hydroxide solution and oxidant into the vanadium precipitate for reaction, and performing solid-liquid separation after the reaction is finished to obtain filtrate A and filter residue A, wherein the weight ratio of vanadium element to phosphorus element in the filtrate A is controlled to be more than or equal to 600; (4) Adding calcium oxide into the filtrate A for precipitating vanadium, and carrying out solid-liquid separation to obtain filtrate B and filter residue B. The method can effectively recycle vanadium in the calcified vanadium extraction tailings, and can reduce vanadium loss in the vanadium extraction process.
Description
Technical Field
The invention relates to the technical field of vanadium extraction metallurgy, in particular to a method for recovering vanadium from calcified vanadium extraction tailings.
Background
Vanadium slag is a main raw material for producing vanadium oxide, and the current industrialization technology comprises two types of sodium roasting-water leaching vanadium extraction and calcification roasting-acid leaching vanadium extraction. The calcified clinker is leached by dilute sulfuric acid under the condition of pH=2.8-3.0, so that vanadium is dissolved into solution, and further ammonium salt is used for precipitating vanadium to prepare vanadium oxide. In the actual production process, the vanadium content of the vanadium extraction tailings generated by the process is 1.2% -1.7%, which is far higher than that of laboratory leaching residues, on the one hand, the leaching pH control range is narrow, when the leaching pH is higher, the calcium pyrovanadate and the calcium manganese pyrovanadate which are supposed to be leached are not fully dissolved, when the leaching pH is lower, vanadium is easy to hydrolyze and precipitate into tailings, and the pH control is completely dependent on a pH detection device, and the device frequently switches between the states of acidic and alkaline slurry, soaking and bare air, and the states of inaccurate measurement such as zero drift and faults are easy to occur in the flushing of the slurry in the use process; on the other hand, the leaching temperature affects the leaching effect, the dissolution speed of the vanadium-containing phase is low when the temperature is low, the leaching rate of vanadium is reduced, vanadium is easy to hydrolyze and precipitate into tailings when the temperature is high, and the temperature of the leaching slurry is affected by the clinker temperature, the leaching agent temperature, the acid temperature and the acid concentration (exothermic in the acid leaching reaction process), the climate temperature and the like. The calcified vanadium extraction tailings have high vanadium content, and the incorrect treatment can cause the waste of vanadium resources.
Vanadium in tailings can be divided into three types: the first is vanadium in the form of calcium pyrovanadate and calcium manganese pyrovanadate, which should be leached but not leached at leaching ph=2.8 to 3.0; the second is distributed in ferric oxide solid solution, pseudobrookite and silicate phase, exists in a wrapping and embedding form, is difficult to leach under the normal condition of pH=2.8-3.0, but can leach in the range of pH=0.5-1.2; and thirdly, the vanadium is hydrolyzed and precipitated into the tailings after being dissolved out due to improper temperature and pH control in the leaching process, and part of vanadium cannot be redissolved into solution by direct acid leaching under the conventional leaching condition.
Aiming at the problem of recovering vanadium from calcified vanadium extraction tailings:
patent application CN 110387468A discloses a method for controlling the secondary acid leaching pH stabilization of calcified clinker, the main technical idea is that the calcified clinker adopts two-stage leaching, the primary leaching uses the calcified clinker and the secondary leaching liquid as raw materials, and the leaching ph=2.5-3.0 is controlled; the secondary leaching takes primary leaching tailings and reuse water as raw materials, and leaching pH=1.5-2.5 is controlled, so that the content of the obtained tailings TV is 1.0% -1.1%. The method has high secondary leaching pH, only can recover a small part of vanadium in tailings, and has low vanadium recovery rate.
Patent application CN 111394576A discloses a method for deep leaching of acid leaching vanadium tailings and solution circulation, and the main technical idea is that calcified clinker adopts two-stage leaching, first-stage leaching takes calcified clinker and residue washing water as raw materials, and leaching pH=2.5-3.2 is controlled; the secondary leaching takes primary leaching tailings, reuse water and part of secondary leaching liquid as raw materials, leaching pH=0.5-1.8 is controlled, and part of the obtained secondary leaching liquid is used for circularly leaching the tailings, and the pH value is regulated to be more than or equal to 2.5 by the reuse water and then is used for washing the primary leaching residues. The method cannot recover vanadium which is precipitated into tailings due to vanadium hydrolysis in the leaching process; the secondary leaching solution is easy to generate hydrolytic precipitation when the vanadium concentration is higher; the secondary leaching solution can cause partial vanadium loss in the process of adjusting the pH value to be more than or equal to 2.5 by using recycled water.
The patent CN 109321760B discloses a recycling method of calcified vanadium extraction tailings, and the main technical idea is that the calcified vanadium extraction tailings are leached by using 2% -2.5% dilute sulfuric acid, the pH value of leaching liquid is adjusted to be 1-2 by using calcified clinker for primary impurity removal, the pH value of primary impurity removal liquid is adjusted to be 2.5-3.5 by using ammonia water or lime milk for secondary impurity removal, the secondary impurity removal liquid is used for leaching the calcified clinker, and the secondary impurity removal slag is returned to a roasting process. According to the method, the vanadium extraction tailings after calcification leaching with dilute sulfuric acid cannot recover vanadium which enters the tailings due to vanadium hydrolysis and precipitation in the leaching process; when clinker is utilized for removing impurities once, the concentration of vanadium in the slurry is high, and the vanadium is easy to hydrolyze to cause vanadium loss.
The prior literature discloses a technical idea of recovering vanadium by low-pH acid leaching of calcified vanadium extraction tailings, but the problems that the vanadium which is hydrolyzed and precipitated in the tailings cannot be recovered, the low-pH acid leaching vanadium has high concentration and is easy to hydrolyze, the vanadium loss is large in the impurity removal process and the like exist.
Disclosure of Invention
The invention aims to solve the problems that vanadium which is hydrolyzed and precipitated in the tailings can not be recovered, the concentration of low-pH acid leaching vanadium is high, hydrolysis is easy, and the vanadium loss in the impurity removal process is large in the low-pH acid leaching vanadium recovery process of the calcified vanadium extraction tailings in the prior art.
In order to achieve the above object, the present invention provides a method for recovering vanadium from calcified vanadium extraction tailings, the method comprising the steps of:
(1) Adding calcified vanadium extraction tailings and calcium sulfite into water for pulping, adding sulfuric acid for leaching, adjusting the pH value of a system to 2-3 by lime and/or limestone after leaching, and then carrying out solid-liquid separation to obtain residues and leaching liquid;
(2) Adjusting the pH value of the leaching solution to 5.5-7 by using magnesium carbonate and/or manganese carbonate to precipitate vanadium, and then filtering to obtain vanadium precipitate and wastewater;
(3) Adding sodium hydroxide solution and oxidant into the vanadium precipitate for reaction, and performing solid-liquid separation after the reaction is finished to obtain filtrate A and filter residue A, wherein the weight ratio of vanadium element to phosphorus element in the filtrate A is controlled to be more than or equal to 600;
(4) Adding calcium oxide into the filtrate A for precipitating vanadium, and then carrying out solid-liquid separation to obtain filtrate B and filter residue B;
wherein, the wastewater returns to the step (1) for use after lime neutralization treatment;
and (3) returning the filtrate B to the step (3) for use.
Preferably, in step (1), V in the calcified vanadium extraction tailings 2 O 5 The content of (2) is 1.4-3 wt%;
further preferably, the liquid-solid ratio of the water to the calcified vanadium extraction tailings is 1-2mL/g.
Preferably, in step (1), the pH of the leaching is between 0.5 and 1.2 and the time of the leaching is between 8 and 30 minutes.
Preferably, in the step (1), the ratio of the amount of the calcium sulfite to the substance of the vanadium element in the calcified vanadium extraction tailings is (0.5-0.7): 1.
preferably, in step (3), the NaOH content of the sodium hydroxide solution is 80-120g/L.
Preferably, in step (3), the liquid-solid ratio of the sodium hydroxide solution to the vanadium precipitate is 8-16mL/g.
Preferably, the oxidant is at least one of air, oxygen and hydrogen peroxide.
Preferably, in step (3), the temperature of the reaction is 80-100 ℃ and the time of the reaction is 30-150min.
Further preferably, in the step (3), the content of tetravalent vanadium in the filtrate A is not more than 0.1g/L.
Preferably, in step (4), the amount of the substance of calcium oxide is 1.3 to 1.6 times the theoretical amount.
Preferably, in the step (4), the temperature of the vanadium precipitation is 90-100 ℃, and the time of the vanadium precipitation is 30-120min.
According to the method, the calcified vanadium extraction tailings are subjected to a reduction acid leaching mode, so that the vanadium which is hydrolyzed and precipitated in the tailings can be recovered, and meanwhile, the problem of vanadium loss caused by the hydrolysis and precipitation of pentavalent vanadium in the low-pH acid leaching process is avoided.
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 recovering vanadium from calcified vanadium extraction tailings, which comprises the following steps:
(1) Adding calcified vanadium extraction tailings and calcium sulfite into water for pulping, adding sulfuric acid for leaching, adjusting the pH value of a system to 2-3 by lime and/or limestone after leaching, and then carrying out solid-liquid separation to obtain residues and leaching liquid;
(2) Adjusting the pH value of the leaching solution to 5.5-7 by using magnesium carbonate and/or manganese carbonate to precipitate vanadium, and then filtering to obtain vanadium precipitate and wastewater;
(3) Adding sodium hydroxide solution and oxidant into the vanadium precipitate for reaction, and performing solid-liquid separation after the reaction is finished to obtain filtrate A and filter residue A, wherein the weight ratio of vanadium element to phosphorus element in the filtrate A is controlled to be more than or equal to 600;
(4) Adding calcium oxide into the filtrate A for precipitating vanadium, and then carrying out solid-liquid separation to obtain filtrate B and filter residue B;
wherein, the wastewater returns to the step (1) for use after lime neutralization treatment;
and (3) returning the filtrate B to the step (3) for use.
In the invention, in the step (1), the calcified vanadium extraction tailings are vanadium slag which is prepared by calcified roasting-acidThe product obtained after the vanadium leaching process is treated, and the V in the calcified vanadium extraction tailings 2 O 5 The content of (C) is 1.4-3 wt%. Specifically, V in the calcified vanadium extraction tailings 2 O 5 The content of (c) may be 1.4 wt%, 1.6 wt%, 1.8 wt%, 2 wt%, 2.2 wt%, 2.4 wt%, 2.6 wt%, 2.8 wt% or 3 wt%.
In the invention, in the step (1), the liquid-solid ratio of the water to the calcified vanadium extraction tailings is 1-2mL/g. Specifically, the liquid-solid ratio of the water to the calcified vanadium extraction tailings may be 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 the invention, in the step (1), the pH value of leaching is 0.5-1.2, and the leaching time is 8-30min. Specifically, the pH of the leaching may be 0.5, 0.6, 0.7, 0.8, 0.9, 1, 1.1 or 1.2, and the leaching time may be 8min, 10min, 12min, 14min, 16min, 18min, 20min, 22min, 24min, 26min, 28min or 30min.
In the present invention, in the step (1), the leaching temperature is an ordinary temperature, and the ordinary temperature is 20-30 ℃. Specifically, the normal temperature may be 20 ℃, 25 ℃ or 30 ℃.
In the invention, in the step (1), the ratio of the amount of the calcium sulfite to the substance of the vanadium element in the calcified vanadium extraction tailings is (0.5-0.7): 1. specifically, the ratio of the amount of calcium sulfite to the substance of vanadium element in the calcified vanadium extraction tailings may be 0.5: 1. 0.55: 1. 0.6: 1. 0.65:1 or 0.7:1.
in particular embodiments, in step (1), the pH of the system may be adjusted to 2, 2.1, 2.2, 2.3, 2.4, 2.5, 2.6, 2.7, 2.8, 2.9 or 3 using lime and/or limestone.
In particular cases, in step (2), the pH of the leachate may be adjusted to 5.5, 5.6, 5.7, 5.8, 5.9, 6, 6.1, 6.2, 6.3, 6.4, 6.5, 6.6, 6.7, 6.8, 6.9 or 7.
In the present invention, in the step (3), the NaOH content in the sodium hydroxide solution is 80-120g/L. Specifically, the NaOH content in the sodium hydroxide solution may be 80g/L, 90g/L, 100g/L, 110g/L or 120g/L.
In the present invention, in the step (3), the liquid-solid ratio of the sodium hydroxide solution to the vanadium precipitate is 8 to 16mL/g. Specifically, the liquid-to-solid ratio of the sodium hydroxide solution to the vanadium precipitate may be 8mL/g, 9mL/g, 10mL/g, 11mL/g, 12mL/g, 13mL/g, 14mL/g, 15mL/g, or 16mL/g.
In the invention, in the step (3), the tetravalent vanadium is oxidized into pentavalent vanadium by adding an oxidant, so that the content of tetravalent vanadium in the filtrate A is low, and the content of tetravalent vanadium in the filtrate A is less than or equal to 0.1g/L in a preferred case.
In the present invention, in the step (3), the oxidizing agent is at least one of air, oxygen and hydrogen peroxide.
In a preferred embodiment, in step (3), the temperature of the reaction is 80-100 ℃ and the time of the reaction is 30-150min. Specifically, the temperature of the reaction may be 80 ℃, 85 ℃, 90 ℃, 95 ℃, or 100 ℃, and the time of the reaction may be 30min, 40min, 50min, 60min, 70min, 80min, 90min, 100min, 110min, 120min, 130min, 140min, or 150min.
In the invention, in the step (3), when the weight ratio of the vanadium element to the phosphorus element in the filtrate A is less than 600, the phosphorus removal agent is added to remove phosphorus, so that the weight ratio of the vanadium element to the phosphorus element in the filtrate A is more than or equal to 600, and when the weight ratio of the vanadium element to the phosphorus element in the filtrate A is more than or equal to 600, the phosphorus removal agent is not added.
Preferably, the dephosphorizing agent is zirconium sulfate and/or calcium oxide.
In the present invention, in the step (4), the amount of the substance of calcium oxide is 1.3 to 1.6 times the theoretical amount. Specifically, the amount of the substance of calcium oxide may be 1.3 times, 1.4 times, 1.5 times, or 1.6 times the theoretical amount.
In the present invention, the theoretical amount refers to the amount of a substance of calcium oxide required to react all vanadium in the filtrate a to produce calcium orthovanadate.
In the preferred case, in the step (4), the temperature of the vanadium precipitation is 90-100 ℃, and the time of the vanadium precipitation is 30-120min. Specifically, the temperature of the vanadium precipitation may be 90 ℃, 92 ℃, 94 ℃, 96 ℃, 98 ℃ or 100 ℃, and the time of the vanadium precipitation may be 30min, 40min, 50min, 60min, 70min, 80min, 90min, 100min, 110min or 120min.
In the invention, the wastewater obtained in the step (2) can be neutralized by lime until the pH value is 8-10, and after removing partial impurity ions such as manganese, magnesium and the like, the wastewater is returned to the step (1) to pulp the calcified vanadium extraction tailings.
In the invention, the sodium hydroxide content in the filtrate B obtained in the step (4) is higher, so that the filtrate B can be returned to the step (3) to react with vanadium precipitate, the partial cyclic utilization of sodium hydroxide is realized, the main chemical component in the filter residue B is calcium vanadate, the vanadium content is higher, and the filtrate B can be returned to the acid leaching vanadium extraction process to extract vanadium together with calcified clinker.
In the step (1), calcium sulfite is selected to reduce vanadium, so that on one hand, the solution stability can be improved, and the influence of pentavalent vanadium hydrolysis precipitation on the vanadium leaching rate is avoided; on the other hand, the oxidation product of the calcium sulfite is calcium sulfate, which is slightly soluble and does not influence the recycling of the subsequent wastewater. After the reaction is finished, lime and/or limestone are/is selected to adjust the pH value of the system to 2-3, so that part of Fe can be removed 3+ Impurities such as P, and the like, tetravalent vanadium is soluble in the pH range, and vanadium loss is very small.
In the step (2), magnesium carbonate and/or manganese carbonate are/is selected to adjust the pH value of the leaching solution to 5.5-7 for precipitating vanadium, on one hand, the speed of adjusting the pH value by carbonate to provide hydroxide radical is low, and bubbles are stirred in the reaction process, so that the obtained vanadium precipitate is precipitated and has good filtering performance; on the other hand, magnesium and manganese ions can not form sediment in the vanadium precipitation process and enter vanadium sediment, and can be separated from a solution system during the neutralization treatment of the wastewater lime, so that the recycling of the wastewater is not influenced. In addition, the carbonate facilitates the control of regulating the pH value to 5.5-7, which is favorable for vanadium and Fe 2+ 、Mn 2+ 、Mg 2+ Separation of the plasma.
In the step (3), the vanadium precipitate in the step (2) is converted into sodium vanadate under alkaline and oxidation conditions, and the alkaline conditions create conditions for dephosphorization, so that dephosphorization or non-dephosphorization can be selected according to the requirements. The oxidant is at least one of air, oxygen and hydrogen peroxide, so that the influence of the introduction of impurity elements on the recycling of the filtrate B is avoided.
The beneficial effects of the invention are as follows:
(1) The calcified vanadium extraction tailings adopt a reduction acid leaching mode, so that the vanadium hydrolyzed and precipitated in the tailings can be recovered, and meanwhile, the problem of vanadium loss caused by the hydrolysis and precipitation of pentavalent vanadium in the low-pH acid leaching process is avoided.
(2) Vanadium in the leaching solution exists in a tetravalent form, so that the problem of large vanadium loss in the process of adjusting the pH and removing impurities is solved.
(3) The obtained filter residue B is returned to the calcified clinker acid leaching process for use, and the subsequent process for independently recovering vanadium is omitted.
(4) The process water can be partially recycled, and impurity elements in the solution are not enriched, so that the main flow of the calcification roasting-acid leaching vanadium extraction process is not influenced.
(5) Vanadium in the calcified vanadium extraction tailings can be fully recovered, and the introduction and enrichment of impurity elements are avoided. The solution system is stable and has good application prospect.
The present invention will be described in detail by way of examples, but the method of the present invention is not limited thereto.
The calcified vanadium extraction tailings used in the following examples and comparative examples are products obtained by treating vanadium slag by a calcified roasting-acid leaching vanadium process, and the main components thereof are shown in table 1.
TABLE 1 main chemical composition/wt% of calcified vanadium extraction tailings
| Name of the name | V 2 O 5 | CaO | MgO | MnO | TiO 2 | SiO 2 | Al 2 O 3 | TFe | Cr 2 O 3 | S |
| No. 1 tailings | 2.88 | 10.15 | 2.56 | 3.74 | 9.98 | 14.47 | 3.14 | 29.70 | 2.05 | 4.45 |
| No. 2 tailings | 2.43 | 7.41 | 2.97 | 3.86 | 10.32 | 14.35 | 3.02 | 29.43 | 1.94 | 3.86 |
| 3# tailings | 2.97 | 9.72 | 2.48 | 3.78 | 9.83 | 14.72 | 3.07 | 29.77 | 1.90 | 4.33 |
Example 1
(1) 1000g of 1 in Table 1 # Adding calcium sulfite and vanadium extraction tailings into 1500mL of water for pulping, adding sulfuric acid for leaching, wherein the leaching pH value is 0.8, the leaching time is 10min, the leaching temperature is 25 ℃, lime is used for adjusting the pH value of a system to 2.5 after the leaching is finished, and then solid-liquid separation is carried out to obtain 1047g of residues and leaching liquid, wherein V in the residues 2 O 5 The content is 0.95 weight percent, and the vanadium leaching rate is 65.53 percent;
(2) Adjusting the pH value of the leaching solution to 6.8 by using magnesium carbonate for precipitating vanadium, and filtering to obtain vanadium precipitate and wastewater, wherein the concentration of TV in the wastewater is 0.04g/L;
(3) Adding 400mL of sodium hydroxide solution (the content of NaOH is 80 g/L) into the vanadium precipitate according to the liquid-solid ratio of the sodium hydroxide solution to the vanadium precipitate being 15.6mL/g, introducing air, reacting for 120min at 85 ℃ under the condition of stirring, and carrying out solid-liquid separation after the reaction is finished to obtain 412mL of filtrate A and filter residue A, wherein the filtrate A contains 24.91g/L of vanadium element and 0.026g/L of phosphorus element, and the content of tetravalent vanadium in the filtrate A is 0.03g/L;
(4) Adding 21.97g of calcium oxide into the filtrate A for precipitating vanadium, wherein the amount of substances of the calcium oxide is 1.3 times of the theoretical amount, the temperature of precipitating vanadium is 95 ℃, the time of precipitating vanadium is 60 minutes, and then carrying out solid-liquid separation to obtain filtrate B and 46.08g of filter residue B;
wherein, the wastewater obtained in the step (2) is neutralized by lime until the pH value is 9, and the wastewater returns to the step (1) to leach the calcified vanadium extraction tailings;
and (3) returning the filtrate B obtained in the step (4) to the step (3) to react with vanadium precipitate.
Example 2
(1) 1000g of 2 in Table 1 # Adding calcium sulfite and 19g of vanadium extraction tailings into 1500mL of water for pulping, adding sulfuric acid for leaching, wherein the leaching pH value is 0.9, the leaching time is 20min, the leaching temperature is 25 ℃, lime is used for adjusting the pH value of a system to 2.7 after the leaching is finished, and then solid-liquid separation is carried out to obtain 1043g of residues and leaching liquid, wherein V in the residues 2 O 5 The content is 0.81 weight percent, and the vanadium leaching rate is 65.33 percent;
(2) Adjusting the pH value of the leaching solution to 6.7 by using magnesium carbonate for precipitating vanadium, and filtering to obtain vanadium precipitate and wastewater, wherein the concentration of TV in the wastewater is 0.04g/L;
(3) Adding 300mL of sodium hydroxide solution (the content of NaOH is 100 g/L) into the vanadium precipitate according to the liquid-solid ratio of the sodium hydroxide solution to the vanadium precipitate being 13.5mL/g, introducing oxygen, reacting for 120min at 90 ℃ under the condition of stirring, and carrying out solid-liquid separation after the reaction is finished to obtain 310mL of filtrate A and filter residue A, wherein the filtrate A contains 28.65g/L of vanadium element and 0.029g/L of phosphorus element, and the content of tetravalent vanadium in the filtrate A is 0.04g/L;
(4) Adding 20.48g of calcium oxide into the filtrate A for precipitating vanadium, wherein the amount of substances of the calcium oxide is 1.4 times of the theoretical amount, the temperature of precipitating vanadium is 95 ℃, the time of precipitating vanadium is 60 minutes, and then solid-liquid separation is carried out to obtain filtrate B and 39.23g of filter residue B;
wherein, the wastewater obtained in the step (2) is neutralized by lime until the pH value is 8, and the wastewater returns to the step (1) to leach the calcified vanadium extraction tailings;
and (3) returning the filtrate B obtained in the step (4) to the step (3) to react with vanadium precipitate.
Example 3
(1) 1000g of 3 in Table 1 # Adding calcium sulfite 23.5g and vanadium extraction tailings into 1500mL water for pulping, adding sulfuric acid for leaching, wherein the leaching pH value is 0.6, the leaching time is 10min, the leaching temperature is 25 ℃, lime is used for adjusting the pH value of the system to 2.8 after the leaching is finished, and then solid-liquid separation is carried out to obtain 1055g of residues and leaching liquid, wherein the V in the residues 2 O 5 The content is 0.87 weight percent, and the vanadium leaching rate is 68.89 percent;
(2) Adjusting the pH value of the leaching solution to 6.5 by using magnesium carbonate for precipitating vanadium, and filtering to obtain vanadium precipitate and wastewater, wherein the concentration of TV in the wastewater is 0.06g/L;
(3) Adding 400mL of sodium hydroxide solution (the content of NaOH is 100 g/L) into the vanadium precipitate according to the liquid-solid ratio of the sodium hydroxide solution to the vanadium precipitate being 14mL/g, introducing air, reacting for 120min at 95 ℃ under the condition of stirring, and carrying out solid-liquid separation after the reaction is finished to obtain 417mL of filtrate A and filter residue A, wherein the filtrate A contains 27.46g/L of vanadium element and 0.028g/L of phosphorus element, and the content of tetravalent vanadium in the filtrate A is 0.04g/L;
(4) Adding 24.52g of calcium oxide into the filtrate A for precipitating vanadium, wherein the amount of substances of the calcium oxide is 1.3 times of the theoretical amount, the temperature of precipitating vanadium is 95 ℃, the time of precipitating vanadium is 60 minutes, and then solid-liquid separation is carried out to obtain filtrate B and 49.38g of filter residue B;
wherein, the wastewater obtained in the step (2) is neutralized by lime until the pH value is 10, and the wastewater returns to the step (1) to leach the calcified vanadium extraction tailings;
and (3) returning the filtrate B obtained in the step (4) to the step (3) to react with vanadium precipitate.
Comparative example 1
The method for recovering vanadium from calcified vanadium extraction tailings by adopting the prior art comprises the following specific steps: 1000g of 2 in Table 1 # Pulping calcified vanadium extraction tailings in 1500mL water, adding sulfuric acid for leaching for 20min, wherein the leaching pH value is 1.8, the leaching temperature is 25 ℃, and solid-liquid separation is carried out after the leaching is finished to obtain leaching solution and 986g residues, wherein V in the obtained residues 2 O 5 The content was 1.78 wt%; the leaching liquid returns to the acid leaching vanadium extraction process for washing calcified vanadium extraction tailings.
Test case
The residue obtained in step (1) of examples 1 to 3 was examined for V in the residue obtained in comparative example 1 2 O 5 V in the residue B obtained in examples 1 to 3 was measured 2 O 5 Is contained in the composition.
The results are shown in Table 2
TABLE 2
| Numbering device | Residue V 2 O 5 Content/wt% | V in residue B 2 O 5 Content/wt% |
| Example 1 | 0.95 | 22.24 |
| Example 2 | 0.81 | 21.96 |
| Example 3 | 0.87 | 22.48 |
| Comparative example 1 | 1.78 | —— |
As can be seen from the results in Table 1, the method of the present invention can effectively recover the vanadium in the calcified vanadium extraction tailings, and the V in the residues 2 O 5 The content is low, the vanadium loss can be reduced, and in the embodiment, the leaching solution is further extracted with vanadium to obtain V 2 O 5 The filter residue B with higher content can be directly returned to the acid leaching vanadium extraction process to recover vanadium together with calcified clinker.
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 (8)
1. A method for recovering vanadium from calcified vanadium extraction tailings, the method comprising the steps of:
(1) Adding calcified vanadium extraction tailings and calcium sulfite into water for pulping, adding sulfuric acid for leaching, adjusting the pH value of a system to 2-3 by lime and/or limestone after leaching, and then carrying out solid-liquid separation to obtain residues and leaching liquid;
(2) Adjusting the pH value of the leaching solution to 5.5-7 by using magnesium carbonate and/or manganese carbonate to precipitate vanadium, and then filtering to obtain vanadium precipitate and wastewater;
(3) Adding sodium hydroxide solution and oxidant into the vanadium precipitate for reaction, and performing solid-liquid separation after the reaction is finished to obtain filtrate A and filter residue A, wherein the weight ratio of vanadium element to phosphorus element in the filtrate A is controlled to be more than or equal to 600;
(4) Adding calcium oxide into the filtrate A for precipitating vanadium, and then carrying out solid-liquid separation to obtain filtrate B and filter residue B;
wherein, the wastewater returns to the step (1) for use after lime neutralization treatment;
the filtrate B is returned to the step (3) for use;
in the step (1), V in the calcified vanadium extraction tailings 2 O 5 The content of the calcium sulfite is 1.4 to 3 weight percent, and the ratio of the calcium sulfite to the mass of vanadium element in the calcified vanadium extraction tailings is (0.5 to 0.7): 1, a step of;
in the step (3), the oxidant is at least one of air, oxygen and hydrogen peroxide, and the content of tetravalent vanadium in the filtrate A is less than or equal to 0.1g/L.
2. The method of claim 1, wherein the liquid to solid ratio of the water to the calcified vanadium extraction tailings is 1-2mL/g.
3. The method according to claim 1, wherein in step (1), the pH of the leaching is 0.5-1.2 and the time of the leaching is 8-30min.
4. The method according to claim 1, wherein in step (3), the NaOH content of the sodium hydroxide solution is 80-120g/L.
5. The method according to claim 1 or 4, characterized in that in step (3) the liquid-solid ratio of the sodium hydroxide solution to the vanadium precipitate is 8-16mL/g.
6. The method according to claim 1, wherein in the step (3), the temperature of the reaction is 80-100 ℃ and the time of the reaction is 30-150min.
7. The method according to claim 1, wherein in step (4), the amount of the substance of calcium oxide is 1.3 to 1.6 times the theoretical amount.
8. The method according to claim 1, wherein in the step (4), the temperature of the vanadium precipitation is 90-100 ℃, and the time of the vanadium precipitation is 30-120min.
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