CN114350981A - 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) adding the calcified vanadium extraction tailings and calcium sulfite into water for pulping, then adding sulfuric acid for leaching, adjusting the pH value of the system to 2-3 by using lime and/or limestone after the pulping, and then carrying out solid-liquid separation to obtain residues and leachate; (2) adjusting the pH value of the leachate 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 a sodium hydroxide solution and an oxidant into the vanadium precipitate for reaction, carrying out solid-liquid separation after the reaction is finished to obtain a filtrate A and a filter residue A, and controlling the weight ratio of vanadium element to phosphorus element in the filtrate A to be more than or equal to 600; (4) and adding calcium oxide into the filtrate A to precipitate vanadium, and carrying out solid-liquid separation to obtain filtrate B and filter residue B. The method can effectively recover the vanadium in the calcified vanadium extraction tailings, and can reduce the 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
The vanadium slag is a main raw material for producing vanadium oxide, and the current industrialized process comprises two types, namely sodium roasting-water vanadium extraction and calcification roasting-acid vanadium extraction. The vanadium slag calcification roasting-acid leaching vanadium extraction process is that vanadium slag and limestone are uniformly mixed and then roasted at high temperature, the obtained calcification clinker is leached by using dilute sulfuric acid under the condition that the pH value is 2.8-3.0, vanadium is dissolved into a solution, and then ammonium salt is precipitated 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 percent and is far higher than that of laboratory leaching residues, and the reasons are that on one hand, the control range of leaching pH is narrow, calcium pyrovanadate and calcium manganese pyrovanadate which are supposed to be leached are insufficiently dissolved when the leaching pH is higher, vanadium is easily hydrolyzed and precipitated to enter tailings when the leaching pH is lower, and the pH control completely depends on a pH detection device which is frequently switched between states of acidic and alkaline slurry, soaking and bare air and scour of the slurry in the using process, and the like, so that states of inaccurate measurement such as zero drift, faults and the like are easily generated; on the other hand, the leaching temperature affects the leaching effect, the dissolution speed of vanadium-containing phase is slow when the temperature is low, the leaching rate of vanadium is reduced, vanadium is easy to hydrolyze and precipitate to enter tailings when the temperature is high, and the temperature of the leached slurry is affected by the temperature of clinker, the temperature of leaching agent, the temperature of acid and the concentration of acid (the heat release in the acid leaching reaction process), the climate temperature and other aspects. The content of vanadium in the calcified vanadium extraction tailings is high, and the waste of vanadium resources can be caused by improper treatment.
The vanadium in the tailings can be divided into three types: the first is vanadium which exists in the form of calcium pyrovanadate and calcium pyrovanadate manganese and should be leached without leaching when the leaching pH is 2.8-3.0; the second one is distributed in iron oxide solid solution, pseudobrookite and silicate phase, exists in a wrapping and inlaying mode, is difficult to leach under the normal condition that the pH value is 2.8-3.0, but can leach within the range of 0.5-1.2; and the third is that the vanadium is dissolved out and hydrolyzed to precipitate the vanadium entering the tailings due to improper control of temperature and pH in the leaching process, and the part of the vanadium can not be dissolved into the solution again 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 pH stability of secondary acid leaching of calcified clinker, which adopts the main technical idea that the calcified clinker is leached in two stages, wherein the calcified clinker and secondary leachate are used as raw materials in the primary leaching, and the pH value of the primary leaching is controlled to be 2.5-3.0; the secondary leaching takes primary leaching tailings and reuse water as raw materials, the leaching pH is controlled to be 1.5-2.5, and the content of the obtained tailings TV is 1.0-1.1%. The method has high secondary leaching pH, can only recover a small part of vanadium in the tailings, and has low vanadium recovery rate.
The 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 is leached in two stages, the calcified clinker and residue washing water are used as raw materials in the first stage leaching, and the pH of the leaching is controlled to be 2.5-3.2; the secondary leaching takes primary leaching tailings, reuse water and part of secondary leaching liquid as raw materials, the leaching pH is controlled to be 0.5-1.8, the obtained secondary leaching liquid is used for partially and circularly leaching the tailings, and the secondary leaching liquid is used for washing the primary leaching residue after the pH is adjusted to be more than or equal to 2.5 by using the reuse water. The method can not recover vanadium entering tailings due to vanadium hydrolysis precipitation in the leaching process; hydrolysis precipitation is easy to occur when the vanadium concentration of the secondary leachate is higher; part of vanadium loss can be caused in the process of adjusting the pH value to be more than or equal to 2.5 by using the reuse water.
The patent CN 109321760B discloses a recycling method of calcification vanadium extraction tailings, and the main technical idea is that the calcification vanadium extraction tailings are leached by using 2% -2.5% dilute sulphuric acid, the leachate is subjected to primary impurity removal by using calcification clinker to adjust the pH value to 1-2, the primary impurity removal liquid is subjected to secondary impurity removal by using ammonia water or lime milk to adjust the pH value to 2.5-3.5, the secondary impurity removal liquid is used for leaching the calcification clinker, and the secondary impurity removal slag returns to the roasting process. The method can not recover vanadium entering tailings due to vanadium hydrolysis precipitation in the leaching process by dilute sulphuric acid leaching calcification vanadium extraction tailings; when clinker is used for primary impurity removal, the concentration of vanadium in the slurry is high, and the vanadium is easy to hydrolyze to cause vanadium loss.
The prior document discloses a technical idea for recovering vanadium by low-pH acid leaching of calcified vanadium extraction tailings, but the problems that vanadium precipitated by hydrolysis in the tailings cannot be recovered, the low-pH acid leaching vanadium has high concentration and is easy to hydrolyze, the vanadium loss in the impurity removal process is large and the like exist.
Disclosure of Invention
The invention aims to solve the problems that in the prior art, when vanadium is recovered by low-pH acid leaching of calcification vanadium extraction tailings, vanadium hydrolyzed and precipitated in the tailings cannot be recovered, the concentration of low-pH acid leaching vanadium is high, the vanadium is easy to hydrolyze, the vanadium loss in the impurity removal process is large, and the like.
In order to achieve the aim, the invention provides a method for recovering vanadium from calcified vanadium extraction tailings, which comprises the following steps:
(1) adding the calcified vanadium extraction tailings and calcium sulfite into water for pulping, then adding sulfuric acid for leaching, adjusting the pH value of a system to 2-3 by using lime and/or limestone after leaching is finished, and then carrying out solid-liquid separation to obtain residues and leachate;
(2) adjusting the pH value of the leachate 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 a sodium hydroxide solution and an oxidant into the vanadium precipitate for reaction, carrying out solid-liquid separation after the reaction is finished to obtain a filtrate A and a filter residue A, and controlling the weight ratio of vanadium element to phosphorus element in the filtrate A to be more than or equal to 600;
(4) adding calcium oxide into the filtrate A for vanadium precipitation, and then carrying out solid-liquid separation to obtain filtrate B and filter residue B;
wherein the wastewater is neutralized by lime and then returns to the step (1) for use;
and (4) returning the filtrate B to the step (3) for use.
Preferably, in the step (1), V in the calcified vanadium extraction tailings2O5In an amount of 1.4 to 3% by weight;
further preferably, the liquid-solid ratio of the water to the calcified vanadium extraction tailings is 1-2 mL/g.
Preferably, in the step (1), the leaching pH value is 0.5-1.2, and the leaching time is 8-30 min.
Preferably, in the step (1), the ratio of the calcium sulfite to the amount of the substance of vanadium element in the calcified vanadium extraction tailings is (0.5-0.7): 1.
preferably, in the step (3), the content of NaOH in the sodium hydroxide solution is 80-120 g/L.
Preferably, in the step (3), the liquid-solid ratio of the sodium hydroxide solution to the vanadium precipitate is 8-16 mL/g.
Preferably, the oxidant is at least one of air, oxygen and hydrogen peroxide.
Preferably, in the step (3), the temperature of the reaction is 80-100 ℃, and the time of the reaction is 30-150 min.
Further preferably, in the step (3), the content of tetravalent vanadium in the filtrate A is less than or equal to 0.1 g/L.
Preferably, in the step (4), the amount of the calcium oxide is 1.3 to 1.6 times of 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 30-120 min.
According to the method, the calcified vanadium extraction tailings are subjected to a reduction acid leaching mode, vanadium hydrolyzed and precipitated in the tailings can be recovered, and the problem of vanadium loss caused by pentavalent vanadium hydrolysis and precipitation in the low-pH acid leaching process is solved.
Detailed Description
The following describes in detail specific embodiments of the present invention. It should be understood that the detailed description and specific examples, while indicating the present invention, are given by way of illustration and explanation only, not limitation.
The endpoints of the ranges and any values disclosed herein are not limited to the precise range or value, and such ranges or values should be understood to encompass values close to those ranges or values. For ranges of values, between the endpoints of each of the ranges and the individual points, and between the individual points may be combined with each other to give one or more new ranges of values, and these ranges of values should 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 the calcified vanadium extraction tailings and calcium sulfite into water for pulping, then adding sulfuric acid for leaching, adjusting the pH value of a system to 2-3 by using lime and/or limestone after leaching is finished, and then carrying out solid-liquid separation to obtain residues and leachate;
(2) adjusting the pH value of the leachate 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 a sodium hydroxide solution and an oxidant into the vanadium precipitate for reaction, carrying out solid-liquid separation after the reaction is finished to obtain a filtrate A and a filter residue A, and controlling the weight ratio of vanadium element to phosphorus element in the filtrate A to be more than or equal to 600;
(4) adding calcium oxide into the filtrate A for vanadium precipitation, and then carrying out solid-liquid separation to obtain filtrate B and filter residue B;
wherein the wastewater is neutralized by lime and then returns to the step (1) for use;
and (4) returning the filtrate B to the step (3) for use.
In the invention, in the step (1), the calcification vanadium extraction tailings are products obtained by treating vanadium slag through a calcification roasting-acid leaching vanadium extraction process, and V in the calcification vanadium extraction tailings2O5Is contained in an amount of 1.4 to 3 wt%. Specifically, V in the calcified vanadium extraction tailings2O5May be present in an amount of 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 1-2 mL/g. Specifically, the liquid-solid ratio of the water to the calcification vanadium extraction tailings can 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 2 mL/g.
In the invention, in the step (1), the pH value of the leaching is 0.5-1.2, and the leaching time is 8-30 min. Specifically, the pH value of the leaching may be 0.5, 0.6, 0.7, 0.8, 0.9, 1, 1.1 or 1.2, and the time of the leaching may be 8min, 10min, 12min, 14min, 16min, 18min, 20min, 22min, 24min, 26min, 28min or 30 min.
In the invention, in the step (1), the leaching temperature is normal temperature, and the normal 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 calcium sulfite to the amount of the substance of vanadium element in the calcified vanadium extraction tailings is (0.5-0.7): 1. specifically, the ratio of the amount of the substances of the vanadium element in the calcium sulfite and 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 instances, 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 invention, in the step (3), the content of NaOH in the sodium hydroxide solution is 80-120 g/L. Specifically, the NaOH content of the sodium hydroxide solution can be 80g/L, 90g/L, 100g/L, 110g/L or 120 g/L.
In the invention, in the step (3), the liquid-solid ratio of the sodium hydroxide solution to the vanadium precipitate is 8-16 mL/g. Specifically, the liquid-to-solid ratio of the sodium hydroxide solution to the vanadium precipitate can be 8mL/g, 9mL/g, 10mL/g, 11mL/g, 12mL/g, 13mL/g, 14mL/g, 15mL/g, or 16 mL/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 the tetravalent vanadium in the filtrate A is lower, and preferably, the content of the tetravalent vanadium in the filtrate A is less than or equal to 0.1 g/L.
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 to 100 ℃ and the time of the reaction is 30 to 150 min. Specifically, the reaction temperature may be 80 ℃, 85 ℃, 90 ℃, 95 ℃ or 100 ℃, and the reaction time may be 30min, 40min, 50min, 60min, 70min, 80min, 90min, 100min, 110min, 120min, 130min, 140min or 150 min.
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, a phosphorus removing agent is added for removing phosphorus, so that the weight ratio of the vanadium element to the phosphorus element in the filtrate A is not less than 600, and when the weight ratio of the vanadium element to the phosphorus element in the filtrate A is not less than 600, the phosphorus removing agent is not added.
Preferably, the phosphorus removal agent is zirconium sulfate and/or calcium oxide.
In the present invention, in the step (4), the amount of the 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 the substance of calcium oxide required to react all the vanadium in the filtrate a to form calcium orthovanadate.
Preferably, in the step (4), the temperature of the vanadium precipitation is 90-100 ℃, and the time of the vanadium precipitation is 30-120 min. Specifically, the temperature of the vanadium precipitation can be 90 ℃, 92 ℃, 94 ℃, 96 ℃, 98 ℃ or 100 ℃, and the time of the vanadium precipitation can be 30min, 40min, 50min, 60min, 70min, 80min, 90min, 100min, 110min or 120 min.
In the invention, the wastewater obtained in the step (2) can be neutralized by lime until the pH value is 8-10, and after partial impurity ions such as manganese, magnesium and the like are removed, the calcified vanadium extraction tailings are returned to the step (1) for pulping.
In the invention, the content of sodium hydroxide in the filtrate B obtained in the step (4) is high, so that the filtrate B can be returned to the step (3) to react with the vanadium precipitate, the local cyclic utilization of sodium hydroxide is realized, the main chemical component in the filter residue B is calcium vanadate, the content of vanadium is high, and the filter residue 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 stability of the solution can be improved, and the influence of hydrolysis and precipitation of pentavalent vanadium on the leaching rate of vanadium is avoided; on the other hand, the oxidation product of the calcium sulfite is calcium sulfate which is slightly soluble, and the recycling of subsequent waste water is not influenced. After the reaction is finished, lime and/or limestone are selected to adjust the pH value of the system to 2-3, so that part of Fe can be removed3+P, etc., tetravalent vanadium is soluble in this pH range with minimal vanadium loss.
In the step (2), magnesium carbonate and/or manganese carbonate are/is selected to adjust the pH value of the leachate to 5.5-7 for vanadium precipitation, on one hand, the speed of providing hydroxide radicals by adjusting the pH value by using carbonate is slow, 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 cannot form precipitates to enter vanadium precipitates in the vanadium precipitation process, and can be separated from a solution system during the neutralization treatment of the waste water lime, so that the recycling use of the waste water is not influenced. In addition, the carbonate is convenient for adjusting the pH value to be 5.5-7 and is beneficial to the control of vanadium and Fe2+、Mn2+、Mg2+And (4) separating the plasma.
In the step (3) of the invention, the vanadium precipitate in the step (2) is converted into sodium vanadate under alkaline and oxidizing conditions, and the alkaline condition creates conditions for phosphorus removal, and phosphorus removal or phosphorus removal can be selected according to requirements. The oxidant is at least one of air, oxygen and hydrogen peroxide, so that the introduction of impurity elements is prevented from influencing the recycling of the filtrate B.
The invention has the beneficial effects that:
(1) the calcified vanadium extraction tailings adopt a reduction acid leaching mode, vanadium hydrolyzed and precipitated in the tailings can be recovered, and the problem of vanadium loss caused by pentavalent vanadium hydrolysis and precipitation in the low-pH acid leaching process is solved.
(2) Vanadium in the leaching solution exists in a tetravalent form, so that the problem of large vanadium loss in the process of adjusting pH and removing impurities is solved.
(3) The obtained filter residue B is returned to the acid leaching process of the calcified clinker for use, and the subsequent process of independently recovering vanadium is omitted.
(4) The process water can be locally recycled, impurity elements in the solution are not enriched, and the main flow of the calcification roasting-acid leaching vanadium extraction process is not influenced.
(5) Can fully recover vanadium in the calcified vanadium extraction tailings and avoid the introduction and enrichment of impurity elements. The solution system is stable and has good application prospect.
The present invention will be described in detail below by way of examples, but the method of the present invention is not limited thereto.
The calcification vanadium extraction tailings used in the following examples and comparative examples are products obtained after vanadium slag is treated by a calcification roasting-acid leaching vanadium extraction process, and the main components of the calcification vanadium extraction tailings are shown in table 1.
TABLE 1 main chemical composition/weight% of calcified vanadium extraction tailings
| Name (R) | V2O5 | CaO | MgO | MnO | TiO2 | SiO2 | Al2O3 | TFe | Cr2O3 | S |
| 1# tailings | 2.88 | 10.15 | 2.56 | 3.74 | 9.98 | 14.47 | 3.14 | 29.70 | 2.05 | 4.45 |
| 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 calcified vanadium extraction tailings and 20g of calcium sulfite into 1500mL of water for pulping, then adding sulfuric acid for leaching, wherein the pH value of leaching is 0.8, and the leaching time isThe leaching time is 10min, the leaching temperature is 25 deg.C, the pH value of the system is adjusted to 2.5 with lime after leaching is finished, and then solid-liquid separation is carried out to obtain 1047g of residue and leachate, V in the residue2O5The content is 0.95 percent by weight, and the vanadium leaching rate is 65.53 percent;
(2) adjusting the pH value of the leachate to 6.8 by using magnesium carbonate to precipitate vanadium, and then filtering to obtain a vanadium precipitate and wastewater, wherein the concentration of TV in the wastewater is 0.04 g/L;
(3) adding 400mL of sodium hydroxide solution (the content of NaOH is 80g/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 after the reaction is finished, carrying out solid-liquid separation 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 0.03 g/L;
(4) adding 21.97g of calcium oxide into the filtrate A to carry out vanadium precipitation, wherein the amount of calcium oxide is 1.3 times of the theoretical amount, the vanadium precipitation temperature is 95 ℃, the vanadium precipitation time is 60min, and then carrying out solid-liquid separation to obtain filtrate B and 46.08g of filter residue B;
neutralizing the wastewater obtained in the step (2) with lime until the pH value is 9, and returning to the step (1) to leach the calcified vanadium extraction tailings;
and (4) returning the filtrate B obtained in the step (4) to the step (3) to react with the vanadium precipitate.
Example 2
(1) 1000g of 2 in Table 1#Adding calcified vanadium extraction tailings and 19g of calcium sulfite into 1500mL of water for pulping, then adding sulfuric acid for leaching, wherein the leached pH value is 0.9, the leaching time is 20min, the leaching temperature is 25 ℃, after leaching, the pH value of a system is adjusted to 2.7 by using lime, then carrying out solid-liquid separation to obtain 1043g of residues and leachate, and V in the residues2O5The content is 0.81 weight percent, and the vanadium leaching rate is 65.33 percent;
(2) adjusting the pH value of the leachate to 6.7 by using magnesium carbonate to precipitate vanadium, and then filtering to obtain a vanadium precipitate and wastewater, wherein the concentration of TV in the wastewater is 0.04 g/L;
(3) adding 300mL of sodium hydroxide solution (the content of NaOH is 100g/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 after the reaction is finished, carrying out solid-liquid separation 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 0.04 g/L;
(4) adding 20.48g of calcium oxide into the filtrate A to precipitate vanadium, wherein the amount of calcium oxide is 1.4 times of the theoretical amount, the vanadium precipitation temperature is 95 ℃, the vanadium precipitation time is 60min, and then carrying out solid-liquid separation to obtain filtrate B and 39.23g of filter residue B;
the wastewater obtained in the step (2) is neutralized by lime until the pH value is 8, and the wastewater is returned to the step (1) to leach the calcified vanadium extraction tailings;
and (4) returning the filtrate B obtained in the step (4) to the step (3) to react with the vanadium precipitate.
Example 3
(1) 1000g of 3 in Table 1#Adding calcified vanadium extraction tailings and 23.5g of calcium sulfite into 1500mL of water for pulping, then adding sulfuric acid for leaching, wherein the pH value of leaching is 0.6, the leaching time is 10min, the leaching temperature is 25 ℃, after leaching, the pH value of a system is adjusted to 2.8 by using lime, then carrying out solid-liquid separation to obtain 1055g of residues and leachate, wherein V in the residues2O5The content is 0.87 weight percent, and the vanadium leaching rate is 68.89 percent;
(2) adjusting the pH value of the leachate to 6.5 by using magnesium carbonate to precipitate vanadium, and then filtering to obtain a vanadium precipitate and wastewater, wherein the concentration of TV in the wastewater is 0.06 g/L;
(3) adding 400mL of sodium hydroxide solution (the content of NaOH is 100g/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 after the reaction is finished, carrying out solid-liquid separation 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 0.04 g/L;
(4) adding 24.52g of calcium oxide into the filtrate A to precipitate vanadium, wherein the amount of calcium oxide is 1.3 times of the theoretical amount, the vanadium precipitation temperature is 95 ℃, the vanadium precipitation time is 60min, and then carrying out solid-liquid separation to obtain filtrate B and 49.38g of filter residue B;
neutralizing the wastewater obtained in the step (2) with lime until the pH value is 10, and returning to the step (1) to leach the calcified vanadium extraction tailings;
and (4) returning the filtrate B obtained in the step (4) to the step (3) to react with the vanadium precipitate.
Comparative example 1
The method comprises the following steps of (1) recovering vanadium from calcified vanadium extraction tailings by adopting the prior art: 1000g of 2 in Table 1#Adding the calcified vanadium extraction tailings into 1500mL of water for pulping, adding sulfuric acid for leaching for 20min, wherein the pH value of the leaching is 1.8, the leaching temperature is 25 ℃, performing solid-liquid separation after leaching is finished to obtain a leaching solution and 986g of residues, and obtaining V in the residues2O5The content was 1.78 wt%; the leachate returns to the acid leaching vanadium extraction process for washing the calcified vanadium extraction tailings.
Test example
The residue obtained in step (1) of examples 1 to 3 and the residue obtained in comparative example 1 were examined for V2O5Content of (D), measurement of V in residue B obtained in example 1-32O5The content of (a).
The results are shown in Table 2
TABLE 2
| Numbering | Residue V2O5Content/weight% | V in filter residue B2O5Content/weight% |
| 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 invention can effectively recover the vanadium in the calcified vanadium extraction tailings, and V in the residues2O5The content is low, the vanadium loss can be reduced, and in the embodiment, vanadium is further extracted from the leaching solution to obtain V2O5The filter residue B with higher content can be directly returned to the acid leaching vanadium extraction process to recover vanadium together with the calcified clinker.
The preferred embodiments of the present invention have been described above in detail, but the present invention is not limited thereto. Within the scope of the technical idea of the invention, many simple modifications can be made to the technical solution of the invention, including combinations of various technical features in any other suitable way, and these simple modifications and combinations should also be regarded as the disclosure of the invention, and all fall within the scope of the invention.
Claims (10)
1. A method for recovering vanadium from calcified vanadium extraction tailings is characterized by comprising the following steps:
(1) adding the calcified vanadium extraction tailings and calcium sulfite into water for pulping, then adding sulfuric acid for leaching, adjusting the pH value of a system to 2-3 by using lime and/or limestone after leaching is finished, and then carrying out solid-liquid separation to obtain residues and leachate;
(2) adjusting the pH value of the leachate 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 a sodium hydroxide solution and an oxidant into the vanadium precipitate for reaction, carrying out solid-liquid separation after the reaction is finished to obtain a filtrate A and a filter residue A, and controlling the weight ratio of vanadium element to phosphorus element in the filtrate A to be more than or equal to 600;
(4) adding calcium oxide into the filtrate A for vanadium precipitation, and then carrying out solid-liquid separation to obtain filtrate B and filter residue B;
wherein the wastewater is neutralized by lime and then returns to the step (1) for use;
and (4) returning the filtrate B to the step (3) for use.
2. The method according to claim 1, characterized in that in step (1), V in the calcified vanadium extraction tailings2O5In an amount of 1.4 to 3% by weight;
preferably, the liquid-solid ratio of the water to the calcified vanadium extraction tailings is 1-2 mL/g.
3. The method as claimed in claim 1, wherein in step (1), the leaching has a pH value of 0.5-1.2 and the leaching time is 8-30 min.
4. The method according to claim 1, characterized in that in step (1), the ratio of the amount of the calcium sulfite to the amount of the substance of vanadium element in the calcified vanadium extraction tailings is (0.5-0.7): 1.
5. the method according to claim 1, wherein in step (3), the NaOH content of the sodium hydroxide solution is 80-120 g/L.
6. The method according to claim 1 or 5, wherein in step (3), the liquid-solid ratio of the sodium hydroxide solution to the vanadium precipitate is 8-16 mL/g.
7. The method according to claim 1 or 6, wherein in step (3), the oxidant is at least one of air, oxygen and hydrogen peroxide.
8. The method according to claim 7, wherein in the step (3), the temperature of the reaction is 80-100 ℃, and the time of the reaction is 30-150 min;
preferably, the content of tetravalent vanadium in the filtrate A is less than or equal to 0.1 g/L.
9. 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.
10. 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 30-120 min.
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