CN114959251A - Vanadium slag roasting leaching method - Google Patents
Vanadium slag roasting leaching method Download PDFInfo
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- CN114959251A CN114959251A CN202210586532.2A CN202210586532A CN114959251A CN 114959251 A CN114959251 A CN 114959251A CN 202210586532 A CN202210586532 A CN 202210586532A CN 114959251 A CN114959251 A CN 114959251A
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- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B26/00—Obtaining alkali, alkaline earth metals or magnesium
- C22B26/20—Obtaining alkaline earth metals or magnesium
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Abstract
The invention relates to a vanadium slag roasting leaching method, and belongs to the technical field of hydrometallurgy. The vanadium slag roasting leaching method comprises the following steps: a. mixing vanadium slag with CaCO 3 Mixing well, then adding into a mixture containing O 2 Roasting in the atmosphere of (1) to obtain roasted clinker; b. the leaching method 1: using CO for the roasting clinker 2 Performing carbonation oxygen-enriched leaching while leaching the CO 2 The volume content of (A) is 85-87%, O 2 The volume content of (A) is 13% -15%; or leaching method 2: and leaching the roasted clinker with acid solution under stirring to obtain vanadium-containing acid leaching solution and acid leaching waste slag vanadium. Book (I)The vanadium slag roasting leaching method has the advantages of no material sticking of the leaching tank, low acid consumption, high vanadium leaching efficiency and use of CO 2 The problems of atmosphere, sulfuric acid corrosion and CaSO are solved during leaching 4 The problem of waste residue land occupation stacking is solved, the purposes of cleaning, environmental protection, low carbon and economy for extracting vanadium are realized, and contribution is made to carbon peak reaching and carbon neutralization.
Description
Technical Field
The invention relates to a vanadium slag roasting leaching method, and belongs to the technical field of hydrometallurgy.
Background
The existing industrialized vanadium slag vanadium extraction methods mainly comprise sodium roasting-water leaching and calcification roasting-acid leaching methods. The sodium treatment roasting-water leaching has great environmental protection pressure due to the problem of difficult treatment of waste water and waste residue. The calcified roasting-acid leaching vanadium extraction realizes the recycling of the waste water, but a large amount of CO is generated in the roasting process 2 Corrosion of acid leaching process, CaSO 4 The waste residue has great pollution to the atmosphere and the land environment, so that a cleaner, more environment-friendly, lower-carbon and more economic vanadium extraction method is urgently needed.
CN 110408772A discloses a method for clean vanadium extraction by vanadium slag roasting, which adopts a calcification roasting method to convert vanadium in vanadium slag into calcium vanadate stable at high temperature, utilizes a high-concentration ammonium sulfate and/or ammonium bisulfate solution to leach roasted clinker, mixes most of ammonium metavanadate crystals and tailings to form crystal-containing slag, and then dissolves the crystal-containing slag in hydrothermal solution to obtain vanadium-containing solution, thereby realizing high-efficiency extraction of vanadium. The method has the advantages that the problem of high-temperature vanadium volatilization is solved, the comprehensive recovery rate of vanadium is higher, ammonia gas is not generated in the leaching process, the conversion rate of vanadium is higher at the same concentration and the same leaching temperature compared with other easily-decomposed ammonium salts, and the problems of low ammonium metavanadate solubility and low vanadium extraction rate in an ammonium salt solution can be effectively solved. The ammonia gas generated by calcining the ammonium metavanadate is absorbed by using a sulfuric acid solution subsequently, and the ammonium sulfate is synthesized again, so that the ammonia can be recycled, the whole process is clean, the production cost is greatly reduced, and the method has good economic benefit and application prospect. However, it still produces CaSO 4 Waste residue and consumption of sulfuric acid solution.
The 'vanadium oxide cleaning process' disclosed in patent application CN101412539A is a high ammonia existing in steel climbing based on the traditional sodium method vanadium extraction processThe method is a cleaning process with zero wastewater discharge, which is redesigned from the technical and economic problems of difficult treatment of nitrogen high-sodium salt wastewater, difficult sale of waste sodium sulfate and the like, and various auxiliary materials, large consumption, low resource utilization rate, high production cost and the like. Laboratory, expansion and semi-industrial tests are carried out for years to build the maximum annual production of 20kt V in world scale 2 O 5 The line for the production of the vanadium slag calcified roasting-sulfuric acid leaching solves the problems of serious ring bonding of a large roasting device, unsmooth production, insufficient leaching dynamics of large leaching equipment and low vanadium yield in sequence, realizes continuous production, obviously improves the productivity, the operation rate, the yield, the benefit and the like, and greatly improves the market competitiveness. See in particular patent applications CN103993160A, CN206474044U, CN106399680B, CN105219976A and CN 105177285A.
The patent applications CN206474044U, CN106399680B and CN105219976A provide a reaction tank stirring device and an acid leaching method with a stirring speed of 'first quick and last slow' (the slurry turning speed is more than or equal to 16 times/min and the stirring speed is 97r/min within 10-15 min; the slurry turning speed is 5.5-8 times/min and the stirring speed is 63r/min after 15 min) aiming at the problem that the acid leaching reaction rate is limited due to the poor dynamic condition caused by insufficient stirring strength in the actual calcified clinker leaching production, mainly solve the problems that the stirring strength at the bottom of a leaching tank is insufficient, stirring dead angles exist, coarse particles with the particle size of about 5 percent more than 180 mu m are not reacted sufficiently, realize turbulent motion caused by stirring of the coarse particle clinker and disperse in the slurry to form vertically uniform turning large-cycle mixing, and generate rapid and close contact and reaction migration with sulfuric acid, and improve the leaching effect, the vanadium content of the acid leaching waste slag is reduced. Wherein the vanadium content of the acid leaching waste slag of the patent application CN105219976A is 1.48-1.71%, which is still higher. In order to solve the problems, the method for roasting and acid leaching the vanadium-containing tailings is further used for recovering vanadium in the patent application CN 109182760A; patent application CN105110373A proposes a secondary leaching method of acid leaching waste slag, and the embodiment of the method can reduce the vanadium content of the acid leaching waste slag to 1.16-1.90%; patent application CN109252047A discloses a clinker acid leaching method for different clinkers, which can reduce acid leaching residue to 1.27%.
Patent application CN105177285A discloses a method for controlling pH stability of acid leaching slurry, which mainly adopts standard correction and offset correction to ensure the stability and reliability of a measurement system; patent application CN110387468A discloses a method for stably controlling pH of two-stage leaching, wherein the vanadium content of the primary acid leaching waste residue is high in industrial implementation, and the vanadium content of the secondary acid leaching waste residue can be reduced to 1.0-1.1%.
The above patent promotes the technical progress of acid leaching vanadium extraction of calcified clinker, the acid leaching stirring technology mainly solves the problem of incomplete coarse particle reaction, the pH control technology mainly solves the problem of reliability of a measurement system, but the vanadium content of acid leaching waste slag is difficult to further reduce.
In addition, the calcium roasting-acid leaching clean vanadium extraction process generates wastewater containing a large amount of metal ions such as vanadium, manganese, magnesium and the like, valuable metals in the process of recovering and leaching vanadium are enriched in a system, so that the impurity content of products in the system is higher, and the production and sale requirements are not met. The discharge amount of the vanadium extraction wastewater can reach 40-50 ten thousand meters 3 In one year, vanadium, manganese and magnesium are respectively about 1g/L, 15g/L and 5g/L, and valuable metal resources of 400-500 tons of vanadium, 6000-7500 tons of manganese and 2000-2500 tons of magnesium are wasted and free acid resources are not recycled every year, so that the waste of valuable metal and free acid resources is greatly caused. If the price of vanadium oxide is estimated to be more than 10 ten thousand yuan/ton, the price of manganese sulfate is about 0.5 ten thousand yuan/ton and the price of manganese sulfate is about 0.05 ten thousand yuan/ton according to the current market price, the direct economic value is lost by about 2 million yuan. At present, the treatment modes of vanadium extraction wastewater comprise an evaporative crystallization method and a lime neutralization method, and the treatment method has the series problems of high treatment cost, no resource recovery of metal ions such as vanadium, manganese, magnesium and the like and no environmental protection risk caused by stacking of sodium sulfate waste residues and gypsum residues. If the wastewater is treated by the traditional way, the treatment cost is up to 20 yuan/m 3 Namely 800-1000 ten thousand yuan per year, 5-7 ten thousand tons of sodium sulfate and gypsum waste residues per year can be generated, and 1.1-1.7 ten thousand tons of carbon dioxide are discharged per year, so that resource waste, environmental pollution and ecological damage are caused.
The calcium calcination-acid leaching vanadium extraction process is shown in figure 9. Only vanadium is extracted, and metals such as Mn, Mg and the like are discarded together with waste residues, so that the resource utilization rate of the calcined vanadium slag is greatly reduced; the method comprises the following steps of (1) leaching vanadium from calcined vanadium slag to obtain a large amount of insoluble substances and soluble substances containing V, Mn, Mg and the like, burying the insoluble substances as solid waste, and obtaining ammonium vanadate from the soluble substances by adopting an ammonium sulfate vanadium precipitation process, wherein the soluble substances have high Mn and Mg contents, so that the ammonium vanadate product has high impurity content, needs to consume a large amount of washing water for washing, and has great influence on the quality of subsequent vanadium trioxide, vanadium pentoxide, FeV50 and FeV 80; and then, carrying out solid-liquid separation by a vanadium precipitation process to obtain a supernatant, removing Mn from the supernatant by a lime neutralization method to generate a large amount of solid waste gypsum residues, and returning the neutralized wastewater to acid leaching after acid adjustment. Although recycling of the wastewater is achieved, a large amount of solid waste is generated. Therefore, the existing process exposes the disadvantages of the above process under the condition of relatively urgent current environmental protection situation: 1) the resource utilization rate is low, the vanadium utilization rate is extremely low, and the Mn and Mg utilization rates are zero; 3) a large amount of solid wastes such as insoluble substances, gypsum slag and the like generated need to be transported and buried, and the environment is greatly polluted due to elements such as iron, titanium, vanadium and the like; 4) the calcination waste heat can not be effectively utilized, and is discharged inefficiently, so that the efficiency loss is caused.
In order to avoid the problems of resource waste of vanadium-manganese-magnesium metal and free acid, large carbon dioxide discharge amount, environmental pollution, high treatment cost and the like caused by the mainstream clean vanadium extraction wastewater treatment process, vanadium-manganese-magnesium metal resource and free acid in the vanadium extraction wastewater can be directly recovered, and no wastewater or waste residue is generated, so that the economic and environmental aims of producing ammonium polyvanadate, manganese sulfate and manganese sulfate products by using the vanadium extraction wastewater resource, reducing wastewater, reducing sodium sulfate waste residue and gypsum waste residue, reducing carbon dioxide, saving wastewater treatment cost and the like are fulfilled, a new technology for directly and hierarchically recovering vanadium-manganese-magnesium by calcining vanadium slag is developed, theoretical and technical supports can be provided for the comprehensive utilization and development of the vanadium extraction clean production process and vanadium slag associated resources, and the method has good economic, social and ecological benefits and contributes to carbon peak reaching and carbon neutralization.
Disclosure of Invention
The first purpose of the invention is to provide a novel vanadium slag roasting leaching method.
In order to achieve the first object of the invention, the vanadium slag roasting leaching method comprises the following steps:
a. mixing vanadium slag with CaCO 3 Mixing well, then adding into a mixture containing O 2 Roasting in the atmosphere of (1) to obtain roasted clinker;
b. the leaching method 1: using CO for the roasting clinker 2 Performing carbonation oxygen-enriched leaching while leaching the CO 2 The volume content of (A) is 85-87%, O 2 The volume content of (A) is 13-15%;
or leaching method 2: leaching the roasted clinker with acid solution under the stirring condition to obtain vanadium-containing acid leaching solution and acid leaching waste slag vanadium, wherein the relationship between the concentration of the acid and the time is as follows: the concentration of the acid solution is 35-65 wt.% in 0-10 min, the concentration of the acid solution is less than 35 wt.% in 10-60 min, and the concentration is more than 10 wt.%;
the stirring is alternately carried out by clockwise stirring and anticlockwise stirring; the acid solution is at least one of sulfuric acid, hydrochloric acid, nitric acid or phosphoric acid;
c. when the leaching method 1 is adopted, after leaching is finished, solid-liquid separation is carried out to obtain an intermediate product vanadium solution and CaCO 3 When the leaching method 2 is adopted, the vanadium solution and the acid insoluble substance which are intermediate products are obtained by solid-liquid separation after the leaching is finished.
Clockwise stirring and anticlockwise stirring are carried out in turn and can adopt the rotatable, reaction tank that has mixing system of base, and reaction tank mixing system clockwise stirring is carried out in turn with base anticlockwise rotation. Or the anticlockwise stirring of the reaction tank stirring system and the clockwise rotation of the base can be alternately carried out.
The roasting in the step a can adopt heating roasting of mixed coal gas, coke oven gas, electricity and other heat sources.
Leaching method 1 the leaching reaction principle mainly comprises:
11Ca 2 V 2 O 7 +Mn 2 V 2 O 7 +22CO 2 +[O 2 ]＝22CaCO 3 +2MnV 12 O 31 +[O 2 ]
in a specific embodiment, the roasting in the step a is carried out at 800-850 ℃ for 1.5-3.5 h to obtain roasted clinker; a step aSaid oxygen-containing group 2 In an atmosphere of (C) O 2 The volume content of the gas is preferably 8-32%, and the rest gas in the atmosphere is at least one of nitrogen, inert gas and combustible gas; more preferably, the atmosphere in step a is O 2 The combustible gas and the air are 8-12 in volume ratio: 10-20: 70-80, and mixing.
In a specific embodiment, the combustible gas in the step a is coke oven gas or mixed gas; a step of CaCO 3 Is limestone; the particle size of the vanadium slag is preferably 63-97 mu m, and the particle size of the limestone is preferably 125-200 mu m; the roasting in the step a is preferably carried out at 805-825 ℃; the roasting time is preferably 1.5-1.9 h; b CO used for leaching in the step 2 And O 2 Preferably CO produced by calcination in step a 2 And O 2 。
In one embodiment, step b the leaching process 2: the concentration of the acid solution is 48-65 wt.% in 0-10 min, the concentration of the acid solution is less than 35 wt.% in 10-60 min, and the concentration is more than 27 wt.%; preferably, the addition rate of the acid solution is: and in 0-10 min, maintaining the relationship between the sulfuric acid concentration C in the pickle liquor and the time t as follows:
and (3) maintaining the relation between the sulfuric acid concentration C in the pickle liquor and the time t for 10-60 min as follows:
the addition rate of the acid solution is preferably 0.0015-0.085 m 3 /min。
In one embodiment, step b the leaching process 2: in the leaching process, the ratio of the consumption of the acid solution to the mass of the calcified clinker is 0.6m 3 /5t～1.5m 3 8 t; preferably, the ratio of the consumption of the acid solution to the mass of the calcified clinker in the leaching process is 0.8m 3 /6.5t～1.0m 3 /7t。
In a specific embodiment, the clockwise stirring speed in the step b is 85-100 r/min, and the counterclockwise stirring speed is 10-50 r/min; preferably, the clockwise stirring speed is 97r/min, and the anticlockwise stirring speed is 20 r/min; after the leaching method 1 in the step b is adopted for leaching, the CaCO is adopted in the step c after the first-stage acid leaching 3 The content of vanadium in the solid mainly containing vanadium is preferably 0.6-1.05%; after leaching in the leaching method 2, the content of vanadium in the acid-insoluble substances in the step c after the first-stage acid leaching is 1.0-1.1 wt.%.
And c, adding ammonia water or ammonium salt into the vanadium solution for precipitating vanadium in the step c, wherein the vanadium precipitation principle is as follows:
MnV 12 O 31 +2NH 4 + ＝(NH 4 ) 2 V 12 O 31 ↓+Mn 2+ 。
for example, the first cycle without adding water to the vanadium-precipitating supernatant and the second cycle with the vanadium-precipitating supernatant may be added with H 2 CO 3 And Mn 2+ The supernatant liquid of the precipitated vanadium can promote the carbonation oxygen-enriched leaching reaction.
In one embodiment, the leaching is the CO-containing produced by calcining in step a 2 、O 2 Leaching the tail gas and the roasting clinker by adding water or vanadium precipitation supernatant, wherein the amount of the added water or vanadium precipitation supernatant is 1-5 times, preferably 1.5-2 times of that of the roasting clinker; the vanadium precipitation supernatant is the supernatant obtained after vanadium precipitation of the vanadium solution in the step c, and the leaching time is preferably 60-120 min;
c step as described with CaCO 3 Returning the solid which is the main solid to the step a for recycling.
The second purpose of the invention is to provide a green and energy-saving method for extracting V, Mn and Mg from the calcined vanadium slag.
In order to achieve the second object of the invention, the method comprises the steps of calcining vanadium slag, leaching vanadium and carrying out solid-liquid separation to obtain insoluble substances and soluble substances;
the soluble substance is extracted by an extracting agent E 1 Extracting to obtain oil phase O 1 And an aqueous phase A 1 Said oil phase O 1 Using acid as stripping agent to make stripping phase separation so as to obtain oil phase O 2 And an aqueous phase A 2 Oil ofPhase O 2 As an extractant E 1 Recycling; the aqueous phase A 2 Precipitating vanadium by using a vanadium precipitation agent, and then carrying out solid-liquid separation to obtain liquid L 1 And solid S 1 Said liquid L 1 Used as back extractant for cyclic utilization, solid S 1 Drying to obtain ammonium vanadate;
the aqueous phase A 1 With the use of an extractant E 2 Extracting and phase-splitting to obtain oil phase O 3 And an aqueous phase A 3 Aqueous phase A 3 The vanadium leaching agent is used for leaching vanadium in the step a for recycling; oil phase O 3 Adopting acid as stripping agent to carry out stripping phase separation to obtain oil phase O 4 And an aqueous phase A 4 Oil phase O 4 As an extractant E 2 Recycling; aqueous phase A 4 Crystallizing at above 60 deg.C, and separating solid and liquid to obtain liquid L 2 And solid S 2 Solid S 2 Drying to obtain MnSO 4 ；
Liquid L 2 Recrystallizing at below 40 deg.C, and separating solid and liquid to obtain liquid L 3 And solid S 3 Liquid L 3 Used as back extractant for cyclic utilization, solid S 3 Drying to obtain MgSO 4 ；
Wherein, the extractant E 1 Is at least one of P204, P507, Cyanex272, DNNSA, N1923, TOA, N263, TBP, naphthenic acid or tertiary carbonic acid, and is preferably P204; the extractant E 2 Is at least one of P204, P507, Cyanex272, DNNSA, N1923, TOA, N263, TBP, naphthenic acid or tertiary carbonic acid, and is preferably P204;
the aqueous phase A 4 Preferably above 100 ℃; liquid L 2 Recrystallization preferably at below 25 ℃;
the stripping agent is preferably H 2 SO 4 、HNO 3 、H 3 PO 4 Or HCl, more preferably sulfuric acid.
The vanadium precipitation agent can be the existing conventional vanadium precipitation agent, such as ammonium salt, and the ammonium salt can be (NH) 4 ) 2 SO 4 And NH 4 HCO 3 And the like.
The vanadium solution may be a vanadium solution obtained by other common calcination-vanadium leaching-solid-liquid separation, for example, the method disclosed in CN 2018112858613. The calcination-vanadium leaching-solid-liquid separation method can also be adopted.
In one embodiment, the solubles include V: 15-45 g/L, Mn: 10-20 g/L, Mg: 1-5 g/L; the acid insoluble matter comprises V: 0.6-1.8 wt.%, Mn: 3.5-7.5 wt.%, Mg: 0.3-0.6 wt.%;
preferably, the vanadium slag is subjected to calcination-vanadium leaching by adopting the vanadium slag roasting leaching method, and the soluble substance is the vanadium solution in the step c.
In one embodiment, the extractant E 1 Is P204; the soluble substance adopts an extracting agent E 1 Compared with the O/A ratio, the extraction ratio of the extraction is preferably 1: 1-1: 10, and more preferably 1: 5;
aqueous phase A 1 With the use of an extractant E 2 The extraction phase ratio O/A of the extraction is 2-3; preferably the aqueous phase A 1 With the use of an extractant E 2 The extraction rate of vanadium is 97.50-100%, the extraction rate of manganese is 40.04-89.84%, and the extraction rate of magnesium is 42.69-81.54%; the extraction frequency is preferably 5-6 times;
the concentration of the back extraction agent is preferably 20 to 60 wt.%;
the O/A ratio of the back extraction is preferably 10: 1-1: 1, and more preferably 3: 1;
and (c) preferably, the drying utilizes the calcination residual heat in the step (a).
Has the advantages that:
1. the inventor of the invention researches and finds that: fine particle clinker with the granularity of below 125 mu m is hardened in the leaching tank to different degrees and mainly concentrates on the clinker feed opening and the tank bottom, and meanwhile, the baffle plate is basically wrapped by the material, so that the axial flow of the solution is influenced, the uneven mixing of sulfuric acid, clinker and mother liquor is possibly caused, and the stability of the leaching effect is influenced. The method of the invention perfectly solves the problems and realizes that:
A. the leaching tank does not stick materials, and the service cycle is longer.
B. The acid consumption is low, the pH control stability is improved, the vanadium content of the primary acid leaching waste slag can be reduced to 0.6-1.1%, wherein the vanadium content can be reduced to 0.6-1.05% by adopting the leaching method 1.
2. The roasting method of the invention is combined with the leaching method 1, thereby solving the problem of CO generated by the calcified roasting of the vanadium slag 2 Atmospheric pollution, corrosion by sulfuric acid, CaSO 4 The waste slag occupies a large area to be stacked, and the aim of extracting vanadium cleanly, environmentally friendly, low-carbon and economically is fulfilled. The main points are as follows:
C. CO produced by calcific roasting 2 The method has the advantages of recycling, zero emission, no pollution to the atmosphere, reduction of greenhouse effect and contribution to carbon peak reaching and carbon neutralization.
D. By using CO 2 Leaching to realize zero corrosion of the equipment field.
3. The utilization rate of V, Mn and Mg resources in the calcined vanadium slag is improved, the utilization rate of the V resources is improved by about 80 percent from about 60 percent, the Mn is improved to about 40 percent from zero, and the Mg is improved to about 20 percent from zero.
4. The method for extracting V, Mn and Mg from the calcined vanadium slag in an environment-friendly and energy-saving manner forms 6 major cycles of zero emission of an extracting agent, vanadium leaching agent, zero emission of a back extracting agent, zero emission of waste heat and the like, and meets the current environment-friendly and energy-saving requirements.
Drawings
FIG. 1 is a flow diagram of a calcination leaching process of the present invention;
FIG. 2 is a flow diagram of the calcination leaching process of comparative example 1;
FIG. 3 is a drawing of a homo-reverse phase stirring apparatus of the present invention, in which the stirring system of the reaction tank is stirring clockwise and the tank base is stirring counterclockwise.
FIG. 4 is a sulfuric acid dropping technical scheme and a vanadium leaching model with the pH of acid leaching slurry changing with time according to the invention.
Fig. 5 shows the concentration of the acid solution used for 0-10 min is 65 wt.%, the concentration of the acid solution used for 10-60 min is 10 wt.%, fig. 5 shows a data picture collected by the central control system in the field implementation process in gray, and the dotted line graph is drawn according to the data collected by the system.
FIG. 6 shows the phenomenon of fine particles sticking to the pot.
FIG. 7A is a graph of pH change for 65% sulfuric acid leaching of vanadium; 7B is a graph of the pH change of 10% sulfuric acid leaching vanadium.
Fig. 8 is a process flow diagram of a leached vanadium solution treatment process of the present invention.
Fig. 9 is a flow chart of a prior art treatment process for a leached vanadium solution.
Detailed Description
In order to achieve the first object of the invention, the vanadium slag roasting leaching method comprises the following steps:
a. mixing vanadium slag with CaCO 3 Mixing well, then adding into a mixture containing O 2 Roasting in the atmosphere of (1) to obtain roasted clinker;
b. the leaching method 1: using CO for the roasting clinker 2 Performing carbonation oxygen-enriched leaching while leaching the CO 2 The volume content of (A) is 85% -87%, O 2 The volume content of (A) is 13-15%;
or leaching method 2: leaching the roasted clinker with acid solution under the stirring condition to obtain vanadium-containing acid leaching solution and acid leaching waste slag vanadium, wherein the relationship between the concentration of the acid and the time is as follows: the concentration of the acid solution is 35-65 wt.% in 0-10 min, the concentration of the acid solution is less than 35 wt.% in 10-60 min, and the concentration is more than 10 wt.%;
the stirring is alternately carried out by clockwise stirring and anticlockwise stirring; the acid solution is at least one of sulfuric acid, hydrochloric acid, nitric acid or phosphoric acid;
c. when the leaching method 1 is adopted, after leaching is finished, solid-liquid separation is carried out to obtain an intermediate product vanadium solution and CaCO 3 As the main solid, when the leaching method 2 is adopted, the intermediate product vanadium solution and the acid insoluble substance are obtained by solid-liquid separation after the leaching is finished.
Clockwise stirring and anticlockwise stirring are carried out in turn and can adopt the rotatable, reaction tank that has mixing system of base, and reaction tank mixing system clockwise stirring is carried out in turn with base anticlockwise rotation. Or the anticlockwise stirring of the reaction tank stirring system and the clockwise rotation of the base can be alternately carried out.
The roasting in the step a can adopt heating roasting of mixed coal gas, coke oven gas, electricity and other heat sources.
Leaching method 1 the leaching reaction principle mainly comprises:
11Ca 2 V 2 O 7 +Mn 2 V 2 O 7 +22CO 2 +[O 2 ]＝22CaCO 3 +2MnV 12 O 31 +[O 2 ]
in a specific embodiment, the roasting in the step a is carried out at 800-850 ℃ for 1.5-3.5 h to obtain roasted clinker; a said step containing O 2 In an atmosphere of (C) O 2 The volume content of the gas is preferably 8-32%, and the rest gas in the atmosphere is at least one of nitrogen, inert gas and combustible gas; more preferably, the atmosphere in step a is O 2 The combustible gas and the air are 8-12 in volume ratio: 10-20: 70-80, and mixing.
In a specific embodiment, the combustible gas in the step a is coke oven gas or mixed gas; a step of CaCO 3 Is limestone; the particle size of the vanadium slag is preferably 63-97 mu m, and the particle size of the limestone is preferably 125-200 mu m; the roasting in the step a is preferably carried out at 805-825 ℃; the roasting time is preferably 1.5-1.9 h; b CO used for leaching in the step 2 And O 2 Preferably CO produced by calcination in step a 2 And O 2 。
In one embodiment, step b the leaching process 2: the concentration of the acid solution is 48-65 wt.% in 0-10 min, the concentration of the acid solution is less than 35 wt.% in 10-60 min, and the concentration is more than 27 wt.%; preferably, the addition rate of the acid solution is: and in 0-10 min, maintaining the relationship between the sulfuric acid concentration C in the pickle liquor and the time t as follows:
and (3) maintaining the relation between the sulfuric acid concentration C in the pickle liquor and the time t for 10-60 min as follows:
the addition rate of the acid solution is preferably 0.0015-0.085 m 3 /min。
In one embodimentB, the leaching method 2: in the leaching process, the ratio of the consumption of the acid solution to the mass of the calcified clinker is 0.6m 3 /5t～1.5m 3 8 t; preferably, in the leaching process, the ratio of the consumption of the acid solution to the mass of the calcified clinker is 0.8m 3 /6.5t～1.0m 3 /7t。
In a specific embodiment, the clockwise stirring speed in the step b is 85-100 r/min, and the counterclockwise stirring speed is 10-50 r/min; preferably, the clockwise stirring speed is 97r/min, and the anticlockwise stirring speed is 20 r/min; after the leaching method 1 in the step b is adopted for leaching, the CaCO is adopted in the step c after the first-stage acid leaching 3 The content of vanadium in the solid mainly containing vanadium is preferably 0.6-1.05%; after the leaching method 2, the content of vanadium in the acid-insoluble substance in the step c after the first-stage acid leaching is 1.0-1.1 wt.%.
And c, adding ammonia water or ammonium salt into the vanadium solution for precipitating vanadium in the step c, wherein the vanadium precipitation principle is as follows:
MnV 12 O 31 +2NH 4 + ＝(NH 4 ) 2 V 12 O 31 ↓+Mn 2+ 。
for example, the first cycle without adding water to the vanadium-precipitating supernatant and the second cycle with the vanadium-precipitating supernatant may be added with H 2 CO 3 And Mn 2+ The supernatant liquid of the precipitated vanadium can promote the carbonation oxygen-enriched leaching reaction.
In one embodiment, the leaching is the CO-containing produced by calcining in step a 2 、O 2 Leaching the tail gas and the roasting clinker by adding water or vanadium precipitation supernatant, wherein the amount of the added water or vanadium precipitation supernatant is 1-5 times, preferably 1.5-2 times of that of the roasting clinker; the vanadium precipitation supernatant is the supernatant obtained after vanadium precipitation of the vanadium solution in the step c, and the leaching time is preferably 60-120 min;
c step as described with CaCO 3 Returning the solid which is the main solid to the step a for recycling.
The second purpose of the invention is to provide a method for extracting V, Mn and Mg from the calcined vanadium slag in a green and energy-saving way.
In order to achieve the second object of the invention, the method comprises the steps of calcining vanadium slag, leaching vanadium and carrying out solid-liquid separation to obtain insoluble substances and soluble substances;
the soluble substance is extracted by an extracting agent E 1 Extracting to obtain oil phase O 1 And an aqueous phase A 1 Said oil phase O 1 Using acid as stripping agent to make stripping phase separation so as to obtain oil phase O 2 And an aqueous phase A 2 Oil phase O 2 As an extractant E 1 Recycling; the water phase A 2 Precipitating vanadium by using a vanadium precipitation agent, and then carrying out solid-liquid separation to obtain liquid L 1 And solid S 1 Said liquid L 1 Used as back extractant for cyclic utilization, solid S 1 Drying to obtain ammonium vanadate;
the aqueous phase A 1 With an extractant E 2 Extracting and phase-splitting to obtain oil phase O 3 And an aqueous phase A 3 Aqueous phase A 3 The vanadium leaching agent is used for leaching vanadium in the step a for recycling; oil phase O 3 Adopting acid as stripping agent to carry out stripping phase separation to obtain oil phase O 4 And an aqueous phase A 4 Oil phase O 4 As an extractant E 2 Recycling; aqueous phase A 4 Crystallizing at above 60 deg.C, and performing solid-liquid separation to obtain liquid L 2 And solid S 2 Solid S 2 Drying to obtain MnSO 4 ；
Liquid L 2 Recrystallizing at below 40 deg.C, and separating solid and liquid to obtain liquid L 3 And solid S 3 Liquid L 3 Used as back extraction agent for cyclic utilization, solid S 3 Drying to obtain MgSO 4 ；
Wherein, the extractant E 1 Is at least one of P204, P507, Cyanex272, DNNSA, N1923, TOA, N263, TBP, naphthenic acid or tertiary carbonic acid, and is preferably P204; the extractant E 2 Is at least one of P204, P507, Cyanex272, DNNSA, N1923, TOA, N263, TBP, naphthenic acid or tertiary carbonic acid, and is preferably P204;
the aqueous phase A 4 Preferably above 100 ℃; liquid L 2 Recrystallization preferably at below 25 ℃;
the stripping agent is preferably H 2 SO 4 、HNO 3 、H 3 PO 4 Or HCl, more preferably sulfuric acid.
The vanadium precipitation agent can be the existing conventional vanadium precipitation agent, such as ammonium salt, and the ammonium salt can be (NH) 4 ) 2 SO 4 And NH 4 HCO 3 And the like.
The vanadium solution may be a vanadium solution obtained by other common calcination-vanadium leaching-solid-liquid separation, for example, the method disclosed in CN 2018112858613. The calcination-vanadium leaching-solid-liquid separation method can also be adopted.
In one embodiment, the solubles include V: 15-45 g/L, Mn: 10-20 g/L, Mg: 1-5 g/L; the acid insoluble matter comprises V: 0.6-1.8 wt.%, Mn: 3.5-7.5 wt.%, Mg: 0.3-0.6 wt.%;
preferably, the vanadium slag is subjected to calcination-vanadium leaching by adopting the vanadium slag roasting leaching method, and the soluble substance is the vanadium solution in the step c.
In one embodiment, the extractant E 1 Is P204; the soluble substance adopts an extracting agent E 1 Compared with the O/A ratio, the extraction ratio of the extraction is preferably 1: 1-1: 10, and more preferably 1: 5;
aqueous phase A 1 With the use of an extractant E 2 The extraction ratio of the extraction is 2-3 compared with O/A; preferably the aqueous phase A 1 With the use of an extractant E 2 The extraction rate of vanadium is 97.50-100%, the extraction rate of manganese is 40.04-89.84%, and the extraction rate of magnesium is 42.69-81.54%; the extraction frequency is preferably 5-6 times;
the concentration of the stripping agent is preferably 20-60 wt.%;
the O/A ratio of the back extraction is preferably 10: 1-1: 1, and more preferably 3: 1;
the drying preferably utilizes the calcining residual heat in the step a.
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
As shown in figure 1, vanadium slag is sieved by a 120-mesh sieve, a 120-mesh undersize product is sieved by a 160-mesh sieve, and a 160-mesh undersize product (the granularity is 63-97 mu m accounting for more than 95 percent and 63 mu m accounting for less than 5 percent) and the granularity are takenMixing limestone powder under 80 mesh standard sieve uniformly and then adding the mixture to O 2 Roasting at 850 ℃ for 2.5 hours under the atmosphere of 10 percent of content, 15 percent of coke oven gas content and 75 percent of air content; adding water to the calcined clinker with CO generated by calcination 2 Performing carbonation oxygen-enriched leaching, wherein the water addition amount is 2.5 times of the roasted clinker amount, and controlling CO in the leaching atmosphere 2 In an amount of 86% by volume, O 2 The volume content of the solution is 14 percent, and the solution is stirred and leached for 60 min; after leaching, the intermediate product vanadium solution and CaCO are obtained by solid-liquid separation 3 Predominantly solids, CaCO 3 The solid which is the main part is returned to be used as limestone powder for recycling. CO 2 2 Leaching to realize zero corrosion of the equipment field.
The V concentration in the vanadium liquid reaches 36.31g/L as CaCO 3 The predominant solid residue, V, was 0.95%.
Adding vanadium solution into (NH) 4 ) 2 SO 4 And H 2 SO 4 Precipitating vanadium, carrying out solid-liquid separation, and returning vanadium precipitation supernatant to be used as leachate.
Example 2
As shown in figure 1, vanadium slag and limestone powder are mixed uniformly and then are added with O 2 Roasting at 840 ℃ in an atmosphere with the content of 12 percent, the content of coke oven gas of 18 percent and the content of air of 70 percent to obtain roasted clinker for 2.5 hours; calcining the calcined clinker to produce CO 2 Performing carbonation oxygen-enriched leaching with water 2.5 times of the amount of the roasted material, and controlling CO in the leaching atmosphere 2 Is 87% by volume, O 2 The volume content of (A) is 15%, and stirring and leaching are carried out for 120 min; after leaching, the intermediate product vanadium solution and CaCO are obtained by solid-liquid separation 3 Predominantly solids, CaCO 3 The solid which is the main part is returned to be used as limestone powder for recycling. CO 2 2 Leaching to realize zero corrosion of the equipment field. The V concentration in the vanadium liquid reaches 35.93g/L as CaCO 3 The predominant solid residue V was 1.05%. Adding vanadium solution into (NH) 4 ) 2 SO 4 And H 2 SO 4 Precipitating vanadium, carrying out solid-liquid separation, and returning vanadium precipitation supernatant to be used as leachate.
Example 3
As shown in fig. 1, willMixing vanadium slag and limestone powder uniformly and then adding the mixture into the mixture 2 Roasting for 1.67 hours at 820 ℃ in the atmosphere of 12 percent of content, 15 percent of coke oven gas content and 73 percent of air content to obtain roasted clinker; calcining the calcined clinker to produce CO 2 Performing carbonation oxygen-enriched leaching with water 1.5 times of the amount of the roasted material, and controlling CO in the leaching atmosphere 2 Is 87% by volume, O 2 The volume content of (A) is 15%, and the stirring leaching is carried out for 80 min; after leaching, the intermediate product vanadium solution and CaCO are obtained by solid-liquid separation 3 Predominantly solids, CaCO 3 The solid which is the main part is returned to be used as limestone powder for recycling. CO 2 2 Leaching to realize zero corrosion of the equipment field.
Calcining the clinker to CO 2 After leaching, the V concentration in the vanadium liquid reaches 38.18g/L as CaCO 3 The predominant solid residue, V, was 0.67%. Adding vanadium solution into (NH) 4 ) 2 SO 4 And H 2 SO 4 Precipitating vanadium, carrying out solid-liquid separation, and returning vanadium precipitation supernatant to be used as leachate.
Example 4
The preparation method of the roasted clinker is the same as that of the embodiment 1, the leaching is carried out by adopting a device shown in figure 3, clockwise stirring and anticlockwise stirring are alternately carried out, the clockwise stirring speed is 90r/min, the anticlockwise stirring speed is 15r/min, 6.5t of the calcified clinker is added, and the mixing ratio is 0.075m 3 Dropwise adding 55% sulfuric acid at a flow rate of/min for 10min, and adding water at a flow rate of 0.0015m 3 Sulfuric acid with the concentration of 10% is dripped at the flow rate of/min. Acid leaching is stopped after 60 min. The vanadium content in the acid leaching residue was 1.05%.
The acid consumption is 0.825m 3 And the tank does not stick materials in the reaction process.
Example 5
The preparation method of the roasted clinker is the same as that of the example 1, the leaching is carried out by adopting the device shown in figure 3, clockwise stirring and anticlockwise stirring are alternately carried out, 7.0t of the calcified clinker is added at the clockwise stirring speed of 97r/min and the anticlockwise stirring speed of 20r/min, and the mixing ratio is 0.085m 3 Dropwise adding 65% sulfuric acid at a flow rate of/min for 10min, and adding water at a flow rate of 0.0025m 3 Sulfuric acid with the concentration of 27% is added dropwise at a flow rate of/min. Acid leaching is stopped after 60 min. Acid leachingThe vanadium content in the slag residue was 1.08%.
The acid consumption is 0.975m 3 And the tank does not stick materials in the reaction process.
Example 6
The process flow of the embodiment is shown in fig. 8, a, leaching vanadium slag by conventional roasting acid, roasting at 900 ℃ for 4h, leaching at pH 2.8-3.2 for 60min, calcifying and calcining the vanadium slag, leaching vanadium by using dilute sulfuric acid with mass concentration of 15%, performing solid-liquid separation to obtain insoluble substances and soluble substances, and measuring V25.40 g/L, Mn 11.10g/L and Mg 1.68g/L in the soluble substances; v1.52 wt.%, Mn 4.11 wt.%, Mg 0.58 wt.% in insolubles;
b. extracting soluble substance with P204 to obtain oil phase O 1 And an aqueous phase A 1 The specific ratio of O/A is 1:5, the extraction times are 5 times, the vanadium extraction rate is 100%, the manganese extraction rate is 89.84%, and the magnesium extraction rate is 81.54%. Oil phase O 1 Dilute sulphuric acid with mass concentration of 20% is used as stripping agent to carry out stripping phase separation to obtain oil phase O 2 And an aqueous phase A 2 Oil phase O 2 The extract is used as an extractant P204 for recycling; aqueous phase A 2 By (NH) 4 ) 2 SO 4 And H 2 SO 4 Precipitating vanadium, and performing solid-liquid separation to obtain liquid L 1 And solid S 1 Liquid L 1 Used as back extraction agent for cyclic utilization, solid S 1 Drying the red cake by using the calcining waste heat to obtain ammonium vanadate;
c. aqueous phase A 1 Extracting and phase-separating by using P15 for 5 times to obtain oil phase O 3 And an aqueous phase A 3 The ratio of O/A of the two is 3:1, the extraction rate of Mn and Mg from P15 is shown in Table 1, and the water phase A 3 The vanadium leaching agent is recycled as the step a; oil phase O 3 Dilute sulphuric acid with mass concentration of 20% is adopted for carrying out back extraction phase separation to obtain oil phase O 4 And an aqueous phase A 4 Back extracted O/A3: 1, oil phase O 4 The extract P15 is recycled; aqueous phase A 4 Crystallizing at 100 deg.C, and separating solid and liquid to obtain liquid L 2 And solid S 2 Solid S 2 Drying by utilizing calcination waste heat to obtain MnSO 4 ；
d. Liquid L 2 By crystallization at 25 deg.C, solid-liquid separationObtaining a liquid L 3 And solid S 3 Liquid L 3 Used as back extractant for cyclic utilization, solid S 3 Drying by using calcination waste heat to obtain MgSO 4 。
TABLE 1 extraction rate of Mn, Mg by P15 grade five
| Number of stages of extraction | Mn/％ | Mg/％ |
| 1 | 40.04 | 42.69 |
| 2 | 47.64 | 44.42 |
| 3 | 76.02 | 69.42 |
| 4 | 82.59 | 72.88 |
| 5 | 89.84 | 81.54 |
As can be seen from the example 1, the method for efficiently extracting V, Mn and Mg by adopting the calcined vanadium slag without waste liquid discharge-solid waste decrement gradingThe green energy-saving method can obtain the MnSO or the ammonium vanadate 4 ，MgSO 4 Three qualified products, namely the extracting agents P204 and P15, the acid leaching agent and the back extraction agent H 2 SO 4 、HNO 3 、H 3 PO 4 And HCl and the like and calcination waste heat are recycled, 6 major cycles of extraction agent zero discharge, vanadium leaching agent, stripping agent zero discharge, waste heat zero discharge and the like are formed, zero generation of gypsum slag is realized, and waste liquid discharge-solid waste reduction-waste heat utilization is realized, so that the utilization rate of vanadium, manganese and magnesium resources is improved, and after P15 five-stage extraction, the vanadium extraction rate reaches 100%, the manganese extraction rate reaches 89.84% and the magnesium extraction rate reaches 81.54%.
The V resource utilization rate is improved by about 80 percent from about 60 percent, the Mn resource utilization rate is improved to about 40 percent from zero, and the Mg resource utilization rate is improved to about 20 percent from zero.
Comparative example 1
As shown in figure 2, vanadium slag and limestone powder are mixed uniformly and then are added with O 2 Roasting at 840 ℃ in an atmosphere with the content of 12 percent, the content of coke oven gas of 18 percent and the content of air of 70 percent to obtain roasted clinker for 2.5 hours; leaching the roasted clinker with 65% of dilute sulfuric acid, wherein the addition amount of the dilute sulfuric acid is 3.0 times of the amount of the roasted clinker, and stirring and leaching for 120 min; after leaching, the intermediate product vanadium solution and CaSO are obtained by solid-liquid separation 4 Mainly solid waste. The V concentration in the vanadium liquid reaches 27.58g/L as CaCO 3 The predominant solid residue V was 1.78%.
Comparative example 2
Otherwise the same as in example 4, except that concentrated sulfuric acid having a concentration of 65% was used throughout, the acid consumption was 1.2m 3 . Vanadium content in acid leach residue 1.53 wt.%. The reaction process is as shown in figure 6.
Comparative example 3
Otherwise the same as in example 4, except that dilute sulfuric acid having a concentration of 10% was used throughout, the acid consumption was 1.5m 3 . The vanadium content in the acid leaching residue was 1.69 wt.%. The reaction process is as shown in figure 6.
Comparative example 4
Otherwise the same as in example 4, except that the stirring was carried out by clockwise stirring only, cisThe hour stirring speed is 90r/min, the acid consumption is 1.3m 3 . The vanadium content in the acid leaching residue was 1.71 wt.%. The fine material is adhered to the bottom and cannot be carried away.
Claims (10)
1. The vanadium slag roasting leaching method is characterized by comprising the following steps:
a. mixing vanadium slag with CaCO 3 Mixing well, then adding into a mixture containing O 2 Roasting in the atmosphere of (1) to obtain roasted clinker;
b. the leaching method 1: using CO for the roasting clinker 2 Performing carbonation oxygen-enriched leaching while leaching the CO 2 The volume content of (A) is 85% -87%, O 2 The volume content of (A) is 13-15%;
or leaching method 2: leaching the roasted clinker with acid solution under the stirring condition to obtain vanadium-containing acid leaching solution and acid leaching waste slag vanadium, wherein the relationship between the concentration of the acid and the time is as follows: the concentration of the acid solution is 35-65 wt.% in 0-10 min, the concentration of the acid solution is less than 35 wt.% in 10-60 min, and the concentration is more than 10 wt.%;
the stirring is alternately carried out by clockwise stirring and anticlockwise stirring; the acid solution is at least one of sulfuric acid, hydrochloric acid, nitric acid or phosphoric acid;
c. when the leaching method 1 is adopted, after leaching is finished, solid-liquid separation is carried out to obtain an intermediate product vanadium solution and CaCO 3 As the main solid, when the leaching method 2 is adopted, the intermediate product vanadium solution and the acid insoluble substance are obtained by solid-liquid separation after the leaching is finished.
2. The vanadium slag roasting leaching method according to claim 1, wherein the roasting in the step a is carried out at 800-850 ℃ for 1.5-3.5 hours to obtain roasted clinker; a said step containing O 2 In an atmosphere of (C) O 2 The volume content of the gas is preferably 8-32%, and the rest gas in the atmosphere is at least one of nitrogen, inert gas and combustible gas; more preferably, the atmosphere in step a is O 2 The combustible gas and the air are 8-12 in volume ratio: 10-20: 70-80 by mixing.
3. According toThe vanadium slag roasting and leaching method according to claim 1 or 2, wherein the combustible gas in the step a is coke oven gas or mixed gas; a step of CaCO 3 Is limestone; the particle size of the vanadium slag is preferably 63-97 mu m, and the particle size of the limestone is preferably 125-200 mu m; the roasting in the step a is preferably carried out at 805-825 ℃; the roasting time is preferably 1.5-1.9 h; b CO used for leaching in the step 2 And O 2 Preferably CO produced by calcination in step a 2 And O 2 。
4. The vanadium slag roasting leaching method according to claim 1 or 2, wherein the leaching method 2 in the step b: the concentration of the acid solution is 48-65 wt.% in 0-10 min, the concentration of the acid solution is less than 35 wt.% in 10-60 min, and the concentration is more than 27 wt.%; preferably, the addition rate of the acid solution is: and in 0-10 min, maintaining the relationship between the sulfuric acid concentration C in the pickle liquor and the time t as follows:
and (3) maintaining the relationship between the sulfuric acid concentration C in the pickle liquor and the time t for 10-60 min as follows:
the addition rate of the acid solution is preferably 0.0015-0.085 m 3 /min。
5. The vanadium slag roasting leaching method according to claim 1 or 2, wherein the leaching method 2 in the step b: in the leaching process, the ratio of the consumption of the acid solution to the mass of the calcified clinker is 0.6m 3 /5t～1.5m 3 8 t; preferably, the ratio of the consumption of the acid solution to the mass of the calcified clinker in the leaching process is 0.8m 3 /6.5t～1.0m 3 /7t。
6. Root of herbaceous plantThe vanadium slag roasting leaching method according to claim 1 or 2, wherein the clockwise stirring speed in the step b is 85-100 r/min, and the counterclockwise stirring speed is 10-50 r/min; preferably, the clockwise stirring speed is 97r/min, and the anticlockwise stirring speed is 20 r/min; after the leaching method 1 in the step b is adopted for leaching, the CaCO is adopted in the step c after the first-stage acid leaching 3 The content of vanadium in the solid mainly containing vanadium is preferably 0.6-1.05%; after leaching in the leaching method 2, the content of vanadium in the acid-insoluble substances in the step c after the first-stage acid leaching is 1.0-1.1 wt.%.
7. The vanadium slag roasting leaching method according to any one of claims 1 to 6, wherein the leaching is to calcine the CO-containing slag generated in the step a to obtain the CO-containing slag 2 、O 2 Leaching the tail gas and the roasting clinker by adding water or vanadium precipitation supernatant, wherein the amount of the added water or vanadium precipitation supernatant is 1-5 times, preferably 1.5-2 times of that of the roasting clinker; the vanadium precipitation supernatant is the supernatant obtained after vanadium precipitation of the vanadium solution in the step c, and the leaching time is preferably 60-120 min;
c step as described with CaCO 3 Returning the solid which is the main part to the step a for recycling.
8. The method for extracting V, Mn and Mg from the calcined vanadium slag in an environment-friendly and energy-saving manner is characterized by comprising the steps of calcining the vanadium slag, soaking vanadium, and carrying out solid-liquid separation to obtain insoluble substances and soluble substances;
the soluble substance is extracted by an extracting agent E 1 Extracting to obtain oil phase O 1 And an aqueous phase A 1 Said oil phase O 1 Using acid as stripping agent to make stripping phase separation so as to obtain oil phase O 2 And an aqueous phase A 2 Oil phase O 2 As an extractant E 1 Recycling; the aqueous phase A 2 Precipitating vanadium by using a vanadium precipitation agent, and then carrying out solid-liquid separation to obtain liquid L 1 And solid S 1 Said liquid L 1 Used as back extractant for cyclic utilization, solid S 1 Drying to obtain ammonium vanadate;
the aqueous phase A 1 With the use of an extractant E 2 Extracting and phase-splitting to obtain oil phase O 3 And an aqueous phase A 3 Aqueous phase ofA 3 The vanadium leaching agent is used for leaching vanadium in the step a for recycling; oil phase O 3 Adopting acid as stripping agent to carry out stripping phase separation to obtain oil phase O 4 And an aqueous phase A 4 Oil phase O 4 As an extractant E 2 Recycling; aqueous phase A 4 Crystallizing at above 60 deg.C, and separating solid and liquid to obtain liquid L 2 And solid S 2 Solid S 2 Drying to obtain MnSO 4 ；
Liquid L 2 Recrystallizing at below 40 deg.C, and separating solid and liquid to obtain liquid L 3 And solid S 3 Liquid L 3 Used as back extractant for cyclic utilization, solid S 3 Drying to obtain MgSO 4 ；
Wherein, the extractant E 1 Is at least one of P204, P507, Cyanex272, DNNSA, N1923, TOA, N263, TBP, naphthenic acid or tertiary carbonic acid, and is preferably P204; the extractant E 2 Is at least one of P204, P507, Cyanex272, DNNSA, N1923, TOA, N263, TBP, naphthenic acid or tertiary carbonic acid, and is preferably P204;
the aqueous phase A 4 Preferably above 100 ℃; liquid L 2 Recrystallization preferably at below 25 ℃;
the stripping agent is preferably H 2 SO 4 、HNO 3 、H 3 PO 4 Or HCl, more preferably sulfuric acid.
9. The green energy-saving method for extracting V, Mn and Mg from calcined vanadium slag according to claim 8, characterized in that the soluble substances comprise V: 15-45 g/L, Mn: 10-20 g/L, Mg: 1-5 g/L; the acid insoluble matter comprises V: 0.6-1.8 wt.%, Mn: 3.5-7.5 wt.%, Mg: 0.3-0.6 wt.%;
preferably, the vanadium slag is subjected to calcination-vanadium leaching by adopting the vanadium slag roasting leaching method of any one of claims 1 to 7, and the soluble substance is the vanadium solution in the step c.
10. The green energy-saving method for extracting V, Mn and Mg from calcined vanadium slag according to claim 8 or 9, characterized in that the extractant E is 1 Is P204; the soluble substance adopts an extracting agent E 1 Compared with the O/A ratio, the extraction ratio of the extraction is preferably 1: 1-1: 10, and more preferably 1: 5;
aqueous phase A 1 With the use of an extractant E 2 The extraction ratio of the extraction is 2-3 compared with O/A; preferably the aqueous phase A 1 With the use of an extractant E 2 The vanadium extraction rate of the extraction is 97.50-100%, the manganese extraction rate is 40.04-89.84%, and the magnesium extraction rate is 42.69-81.54%; the extraction frequency is preferably 5-6 times;
the concentration of the stripping agent is preferably 20-60 wt.%;
the O/A ratio of the back extraction is preferably 10: 1-1: 1, and more preferably 3: 1;
and (c) preferably, the drying utilizes the calcination residual heat in the step (a).
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| CN110129575A (en) * | 2019-06-24 | 2019-08-16 | 攀钢集团研究院有限公司 | Roasting clinker pressure leaching prepares method of the low sodium vanadium than leachate |
| WO2021155441A1 (en) * | 2020-02-07 | 2021-08-12 | Avanti Materials Ltd | Recovery of vanadium from alkaline slag materials |
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