CN114959251B - Vanadium slag roasting leaching method - Google Patents

Vanadium slag roasting leaching method Download PDF

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CN114959251B
CN114959251B CN202210586532.2A CN202210586532A CN114959251B CN 114959251 B CN114959251 B CN 114959251B CN 202210586532 A CN202210586532 A CN 202210586532A CN 114959251 B CN114959251 B CN 114959251B
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vanadium
leaching
roasting
acid
vanadium slag
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CN114959251A (en
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杨林
王进
周清烈
王宝琦
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Sichuan University
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    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/02—Roasting processes
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • 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
    • C22B26/22—Obtaining magnesium
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B34/00—Obtaining refractory metals
    • C22B34/20—Obtaining niobium, tantalum or vanadium
    • C22B34/22—Obtaining vanadium
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B47/00—Obtaining manganese
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • C22B7/006—Wet processes
    • C22B7/007—Wet processes by acid leaching
    • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00—Technologies related to metal processing
    • Y02P10/20—Recycling

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  • Manufacture And Refinement Of Metals (AREA)

Abstract

The invention relates to a vanadium slag roasting leaching method, and belongs to the technical field of hydrometallurgy. The vanadium slag roasting and leaching method comprises the following steps: a. mixing vanadium slag with CaCO 3 Mixing uniformly and then adding O 2 Roasting in the atmosphere of (2) to obtain roasting clinker; b. leaching method 1: CO for the roasted clinker 2 Performing carbonation oxygen-enriched leaching, wherein CO is obtained during leaching 2 The volume content of (2) is 85-87%, O 2 The volume content of (3) is 13-15%; or leaching method 2: and leaching vanadium in the calcified clinker by using an acid solution under the stirring condition to obtain a pickling solution containing vanadium and acid leaching waste slag vanadium. The vanadium slag roasting leaching method has the advantages of no sticking of materials in a leaching tank, low acid consumption, high vanadium leaching efficiency and CO use 2 The leaching also solves the problems of atmospheric, sulfuric acid corrosion and CaSO 4 The waste residue occupies the land and is piled up, thereby realizing the aims of cleaning, environmental protection, low carbon and economic vanadium extraction and contributing to carbon peak and carbon neutralization.

Description

Vanadium slag roasting leaching method
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 method mainly comprises sodium roasting-water leaching and calcified roasting-acid leaching methods. Sodium roasting-water leaching is extremely high in environmental protection pressure due to the problem of difficult treatment of waste water and waste residues. The calcified roasting-acid leaching vanadium realizes the recycling of waste water, but due to a large amount of CO generated in the roasting process 2 Corrosion and CaSO of acid leaching process 4 Waste residues are particularly polluting the atmosphere and the land, so that a cleaner, more environment-friendly, lower-carbon and more economical vanadium extraction method is urgently needed.
CN 110408772A discloses a method for roasting and cleaning vanadium slag to extract vanadium, which adopts calcified roasting method to convert vanadium in the vanadium slagThe method is characterized in that calcium vanadate stable at high temperature is obtained, high-concentration ammonium sulfate and/or ammonium bisulfate solution is used for leaching roasting clinker, most of ammonium metavanadate crystals and tailings are mixed to form crystal-containing slag, and then the crystal-containing slag is dissolved in hot liquid to obtain vanadium-containing solution, so that high-efficiency extraction of vanadium is realized. The method has the advantages that the method has no high-temperature vanadium volatilization problem, the comprehensive recovery rate of vanadium is higher, ammonia gas is not generated in the leaching process, compared with other ammonium salts which are easy to decompose, the method has higher conversion rate of vanadium at the same concentration and the same leaching temperature, and meanwhile, the problems of small solubility of ammonium metavanadate in the ammonium salt solution and low extraction rate of vanadium can be effectively avoided. The ammonia generated by calcining ammonium metavanadate is absorbed by the sulfuric acid solution subsequently, so that ammonium sulfate is synthesized again, 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 consume sulfuric acid solution.
The 'vanadium oxide cleaning process' disclosed in the patent application CN101412539A is an environment-friendly difficult problem that high ammonia nitrogen and high sodium salt wastewater is difficult to treat, waste sodium sulfate is difficult to sell and the like in the traditional sodium method vanadium extraction process of climbing steel, and the technology and economic difficult problems that auxiliary materials are various, the consumption is large, the resource utilization rate is low, the production cost is high and the like, and the redesigned wastewater is free from emission. Laboratory, expansion and semi-industrial tests are carried out for years, and the annual yield of 20kt V with the maximum world model is built 2 O 5 The vanadium slag calcification roasting-sulfuric acid leaching production line sequentially solves the problems of serious kiln sticking, unsmooth production, insufficient leaching dynamics of large-scale leaching equipment and low vanadium yield of large-scale roasting equipment, realizes continuous production, remarkably improves productivity, operation rate, yield, benefit and the like, and greatly improves market competitiveness. See in particular patent applications CN103993160A, CN206474044U, CN106399680B, CN105219976a and CN105177285A.
The patent applications CN206474044U, CN106399680B and CN105219976A provide a reaction tank stirring device and a stirring speed of 'fast and slow first' (within 10-15 min, the slurry turning speed is more than or equal to 16 times/min, the stirring speed is 97r/min; after 15min, the slurry turning speed is 5.5-8 times/min, and the stirring speed is 63 r/min), aiming at solving the problems that the stirring strength at the bottom of a leaching tank is insufficient, the stirring dead angle exists, and coarse particles with the granularity of more than 180 mu m react insufficiently, the coarse particles are involved in the stirring to cause turbulent motion and dispersed in the slurry to form large circulation mixing which turns up and down uniformly, the rapid close contact and reaction migration with sulfuric acid occur, the leaching effect is improved, and the purpose of reducing the vanadium content of acid leaching waste slag is mainly solved. Wherein the content of vanadium in the acid leaching waste slag of the patent application CN105219976A is 1.48-1.71 percent, and the content is still higher. In order to solve the problems, the patent application CN109182760A further recovers vanadium by a method of roasting and acid leaching vanadium-containing tailings; patent application CN105110373A proposes a method for secondary leaching of acid leaching waste slag, and the vanadium content of the acid leaching waste slag can be reduced to 1.16-1.90% from the embodiment; patent application CN109252047a discloses a clinker acid leaching method for different clinker, which can reduce the acid leaching residue to 1.27%.
Patent application CN105177285a discloses a method for controlling pH stability of acid leaching slurry, mainly adopting standard correction and offset correction, ensuring stability and reliability of a measurement system; patent application CN110387468A discloses a method for controlling the pH stability of two-stage leaching, wherein the vanadium content of the primary acid leaching waste slag is high in the industrialized implementation, and the vanadium content of the secondary acid leaching waste slag can be reduced to 1.0-1.1%.
The above patent promotes the technical progress of calcified clinker acid leaching vanadium extraction, the acid leaching stirring technology mainly solves the problem of incomplete coarse particle reaction, and the pH control technology mainly solves the problem of reliability of a measuring system, but further reduces the vanadium content of acid leaching waste slag.
In addition, the calcium roasting-acid leaching clean vanadium extraction process generates wastewater containing a large amount of vanadium, manganese, magnesium and other metal ions, valuable metals in the vanadium leaching recovery process are enriched in the system, so that the impurity content of products in the system is higher, and the production and sales requirements are not met. The discharge amount of the vanadium extraction wastewater can reach 40-50 ten thousand meters 3 The vanadium, manganese and magnesium are contained in the steel at the concentration of about 1g/L, 15g/L and 5g/L respectively, and 400-500 tons of vanadium and 6 tons of vanadium are contained in the steel every yearThe valuable metal resources of 000-7500 tons of manganese and 2000-2500 tons of magnesium are wasted and the free acid resources are not recovered, so that the great valuable metal and free acid resources are wasted. If the price of vanadium oxide is estimated according to the current market price, 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, so that the direct economic value loss is about 2 hundred million yuan. The existing treatment modes of vanadium extraction wastewater comprise an evaporation crystallization method and a lime neutralization method, and have the series of 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 piling up of sodium sulfate waste residues and gypsum residues. If the wastewater is treated in the traditional way, the treatment cost is up to 20 yuan/m 3 Namely 800-1000 ten thousand yuan/year, 5-7 ten thousand tons/year of sodium sulfate and gypsum waste residues can be produced, and 1.1-1.7 ten thousand tons/year of carbon dioxide is discharged, so that resource waste, environmental pollution and ecological damage are caused.
The calcium calcination-acid leaching vanadium extraction process is shown in fig. 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 vanadium slag is calcined to obtain a large amount of insoluble matters and soluble matters containing V, mn, mg and the like after vanadium leaching, the insoluble matters are used as solid waste for landfill treatment, the soluble matters are subjected to an ammonium sulfate vanadium precipitation process to obtain ammonium vanadate, and the impurity content of an ammonium vanadate product is high due to the high content of the soluble matters Mn and Mg, a large amount of washing water is required to be consumed for washing, and the quality influence on the subsequent vanadium trioxide, vanadium pentoxide, feV50 and FeV80 is large; and in addition, supernatant is obtained through solid-liquid separation of a vanadium precipitation process, mn is removed from the supernatant through a lime neutralization method to generate a large amount of solid waste gypsum slag, and the neutralized wastewater is subjected to acid regulation and then returned to acid leaching. Although recycling of the waste water is achieved, a large amount of solid waste is produced. Therefore, under the condition that the current environment protection situation is urgent, the prior art exposes a plurality of defects existing in the prior art: 1) The utilization rate of resources is low, the utilization rate of vanadium is extremely low, and the utilization rate of Mn and Mg is zero; 3) The generated large amount of solid wastes such as insoluble matters, gypsum slag and the like need to be transported and landfilled, and the generated large amount of solid wastes contain elements such as iron, titanium, vanadium and the like, so that the environment is greatly polluted; 4) The calcining waste heat is not effectively utilized, and is discharged in an ineffective way, so that the effective efficiency is lost.
In order to avoid the problems of vanadium, manganese and magnesium metal and free acid resource waste, large carbon dioxide emission, high environmental pollution, high treatment cost and the like caused by the main flow clean vanadium extraction wastewater treatment process, the vanadium, manganese and magnesium metal resource and free acid in the vanadium extraction wastewater can be directly recovered, no wastewater or waste residue is generated, the production of ammonium polyvanadate, manganese sulfate and manganese sulfate products by utilizing the vanadium extraction wastewater resource, the emission reduction of waste water, sodium sulfate waste residue and gypsum waste residue are realized, the emission reduction of carbon dioxide is realized, the wastewater treatment cost and other economic and environmental purposes are saved, the new technology for directly recovering vanadium, manganese and magnesium by classification by calcining vanadium slag is developed, theoretical and technical support can be provided for the comprehensive utilization and development of the vanadium extraction clean production process and the vanadium slag associated resource, and good economic, social and ecological benefits are contributed to peak carbon neutralization.
Disclosure of Invention
The first object 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 and leaching method comprises the following steps:
a. mixing vanadium slag with CaCO 3 Mixing uniformly and then adding O 2 Roasting in the atmosphere of (2) to obtain roasting clinker;
b. leaching method 1: CO for the roasted clinker 2 Performing carbonation oxygen-enriched leaching, wherein CO is obtained during leaching 2 The volume content of (2) is 85-87%, O 2 The volume content of (3) is 13-15%;
or leaching method 2: leaching vanadium in the calcified clinker by using an acid solution under the stirring condition to obtain a pickling solution containing vanadium and acid leaching waste slag vanadium, wherein the concentration of the acid and the time are as follows: the concentration of the acid solution is 35-65 wt.% within 0-10 min, and the concentration of the acid solution is less than 35wt.% and more than 10wt.% within 10-60 min;
the stirring is performed alternately 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, the intermediate product is obtained by solid-liquid separation after the leaching is finishedVanadium solution and CaCO 3 And when the leaching method 2 is adopted for the main solid, the intermediate product vanadium solution and acid insoluble matters are obtained through solid-liquid separation after the leaching is finished.
The clockwise stirring and the anticlockwise stirring are alternately carried out, and a reaction tank with a rotatable base and a stirring system can be adopted, and the clockwise stirring and the anticlockwise rotation of the base are alternately carried out by the stirring system of the reaction tank. Or the counter-clockwise stirring and the clockwise rotation of the base of the reaction tank stirring system are alternately performed.
The roasting in the step a can be performed by heating with heat sources such as mixed gas, coke oven gas, electricity and the like.
The leaching reaction principle of the leaching method 1 mainly comprises the following steps:
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 for 1.5 to 3.5 hours at 800 to 850 ℃ to obtain roasted clinker; the step a comprises O 2 O in atmosphere of (2) 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 such that O 2 The volume ratio of the combustible gas to the air is 8-12: 10-20: 70-80.
In a specific embodiment, the combustible gas in the step a is coke oven gas or mixed gas; step a, caCO 3 Is limestone; the granularity of the vanadium slag is preferably 63-97 mu m, and the granularity of the limestone is preferably 125-200 mu m; the roasting in the step a is preferably roasting at 805-825 ℃; the roasting time is preferably 1.5-1.9 h; the CO adopted by the leaching in the step b 2 And O 2 Preferably CO from the calcination of step a 2 And O 2 。
In a specific embodiment, the leaching method 2 of step b: the concentration of the acid solution is 48-65 wt.% within 0-10 min, and the concentration of the acid solution is less than 35wt.% and more than 27wt.% within 10-60 min; preferably, the acid solution is added at a rate of: the relation between the sulfuric acid concentration C in the pickle liquor and the time t is maintained within 0-10 min:
maintaining the relation between the sulfuric acid concentration C and the time t in the pickle liquor for 10-60 min:
the acid solution is preferably added at a rate of 0.0015 to 0.085m 3 /min。
In a specific embodiment, the leaching method 2 of 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 8t; preferably, the ratio of the consumption of acid solution to the mass of calcified clinker during the leaching 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 anticlockwise stirring speed is 10-50 r/min; preferably, the clockwise stirring speed is 97r/min, and the anticlockwise stirring speed is 20r/min; after leaching by the leaching method 1 in the step b, the CaCO is used in the step c after primary acid leaching 3 The vanadium content in the main solid is preferably 0.6-1.05%; after leaching in leaching method 2, the content of vanadium in the acid insoluble matters in step c after primary acid leaching is 1.0-1.1 wt.%.
And c, adding ammonia water or ammonium salt into the vanadium solution for precipitating vanadium, wherein the principle of precipitating vanadium is as follows:
MnV 12 O 31 +2NH 4 + ＝(NH 4 ) 2 V 12 O 31 ↓+Mn 2+ 。
for example, the supernatant liquid without vanadium precipitation in the primary circulation is added with water, and the supernatant liquid with vanadium precipitation in the secondary circulation can be added with H 2 CO 3 And Mn of 2+ And the supernatant liquid of the vanadium precipitation can promote the carbonation oxygen-enriched leaching reaction.
In one embodimentWherein the leaching is CO-containing produced by calcining the step a 2 、O 2 Adding water or precipitating vanadium supernatant to leach out the tail gas of the roasting clinker, wherein the amount of the added water or precipitating vanadium supernatant is 1-5 times, preferably 1.5-2 times, of the amount of the roasting clinker; the supernatant fluid after vanadium precipitation of the vanadium solution in the step c is the supernatant fluid after vanadium precipitation, and the leaching time is preferably 60-120 min;
CaCO as described in step c 3 The main solid is returned to the step a for recycling.
The second purpose of the invention is to provide a green energy-saving method for extracting V, mn and Mg from calcined vanadium slag.
In order to achieve the second object of the invention, the method comprises the steps of calcining vanadium slag, immersing vanadium and carrying out solid-liquid separation to obtain insoluble matters and soluble matters;
the soluble substances are extracted by an extractant E 1 Extraction to obtain oil phase O 1 And aqueous phase A 1 The oil phase O 1 Back extraction phase separation is carried out by taking acid as a back extraction agent to obtain oil phase O 2 And aqueous phase A 2 Oil phase O 2 As extractant E 1 Recycling; the aqueous phase A 2 Vanadium is precipitated by adopting a vanadium precipitating agent, and then solid-liquid separation is carried out to obtain liquid L 1 And solids S 1 The liquid L 1 Can be used as back extraction agent for recycling, and solid S 1 Drying to obtain ammonium vanadate;
the aqueous phase A 1 By using extractant E 2 Extracting and phase-separating to obtain oil phase O 3 And aqueous phase A 3 Aqueous phase A 3 Recycling the vanadium leaching agent used as the vanadium leaching agent for the step a; oil phase O 3 Back extraction phase separation is carried out by adopting acid as a back extraction agent to obtain oil phase O 4 And aqueous phase A 4 Oil phase O 4 As extractant E 2 Recycling; aqueous phase A 4 Crystallizing at above 60deg.C, and separating solid from liquid to obtain liquid L 2 And solids S 2 Solids S 2 Drying to obtain MnSO 4 ；
Liquid L 2 Recrystallizing below 40deg.C, and separating solid from liquid to obtain liquid L 3 And solids S 3 Liquid L 3 Can be used as back extraction agent for recycling, and solid S 3 DryingTo obtain MgSO 4 ；
Wherein the extractant E 1 At least one of P204, P507, cyanex272, DNNSA, N1923, TOA, N263, TBP, naphthenic acid or tertiary carbonic acid, preferably P204; the extractant E 2 At least one of P204, P507, cyanex272, DNNSA, N1923, TOA, N263, TBP, naphthenic acid or tertiary carbonic acid, preferably P204;
the aqueous phase A 4 Preferably at above 100 ℃; liquid L 2 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 precipitating agent may be an existing conventional vanadium precipitating agent, for example, an ammonium salt, which may be (NH) 4 ) 2 SO 4 And NH 4 HCO 3 Etc.
The vanadium solution may be a vanadium solution obtained by other common calcination-vanadium leaching-solid-liquid separation methods, such as 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 includes V:0.6 to 1.8wt.%, mn:3.5 to 7.5wt.%, mg:0.3 to 0.6wt.%;
preferably, the vanadium slag is calcined and leached by the vanadium slag roasting leaching method, and the soluble matters are the vanadium solution in the step c.
In one embodiment, the extractant E 1 P204; the soluble substance adopts extractant E 1 The extraction ratio O/A of the extraction is preferably 1:1 to 1:10, more preferably 1:5;
aqueous phase A 1 By using extractant E 2 The extraction ratio O/A of the extraction is 2-3; preferably the aqueous phase A 1 By using 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%; secondary of extractionThe number is preferably 5 to 6 times;
the concentration of the stripping agent is preferably 20-60 wt.%;
the ratio O/A of the back extraction is preferably 10:1-1:1, more preferably 3:1;
the drying preferably utilizes the waste heat of the calcination in the step a.
The beneficial effects are that:
1. the inventors of the present invention studied and found that: the inside of the leaching tank is hardened by fine-grained clinker with granularity less than 125 mu m to different degrees, the fine-grained clinker is mainly concentrated at the clinker discharging opening and the tank bottom part, and meanwhile, the baffle plate is also basically wrapped by materials, 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 the following steps:
A. the leaching tank is not sticky, and the service period is long.
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%, and the vanadium content can be reduced to 0.6-1.05% when the leaching method 1 is adopted.
2. The roasting method of the invention is combined with the leaching method 1, and solves the problem of CO generated by calcified roasting of vanadium slag 2 Contaminated atmosphere, sulfuric acid corrosion, caSO 4 The waste residue occupies the land and is piled up, thereby realizing the aims of cleaning, environmental protection, low carbon and economic vanadium extraction. The main expression is as follows:
C. CO produced by calcification 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 and carbon neutralization.
D. By CO 2 Leaching to realize zero corrosion of equipment sites.
3. The utilization rate of V, mn and Mg resources in the calcined vanadium slag is improved, the utilization rate of V resources is improved by about 80 percent from about 60 percent, mn is improved to about 40 percent from zero, and Mg is improved to about 20 percent from zero.
4. The method for extracting the V, mn and Mg in the calcined vanadium slag in the environment-friendly and energy-saving way forms 6 large cycles of zero discharge of the extractant, zero discharge of the vanadium leaching agent, zero discharge of the back extractant, zero discharge of waste heat and the like, and meets the current environment-friendly and energy-saving requirement.
Drawings
FIG. 1 is a flow chart of a calcination leaching process of the present invention;
FIG. 2 is a flow chart of the calcination leaching process of comparative example 1;
FIG. 3 is a diagram of a co-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 shows a sulfuric acid dripping technical scheme and a time-varying vanadium leaching model of pH of acid leaching slurry.
Fig. 5 shows that the concentration of the acid solution adopted in 0-10 min is 65wt.%, the concentration of the acid solution adopted in 10-60 min is 10wt.%, and gray in fig. 5 is a graph showing data collected by a central control system in the field implementation process, and a line graph in the graph is drawn according to the data collected by the system.
Fig. 6 shows the phenomenon of sticking of fine clinker particles to the tank.
FIG. 7A is a graph showing the pH change of a 65% sulfuric acid vanadium leach; 7B is a graph showing the pH change of 10% sulfuric acid vanadium leaching.
FIG. 8 is a flow chart of a process for treating leached vanadium solution in accordance with the present invention.
Fig. 9 is a flow chart of a prior art process for treating a leached vanadium solution.
Detailed Description
In order to achieve the first object of the invention, the vanadium slag roasting and leaching method comprises the following steps:
a. mixing vanadium slag with CaCO 3 Mixing uniformly and then adding O 2 Roasting in the atmosphere of (2) to obtain roasting clinker;
b. leaching method 1: CO for the roasted clinker 2 Performing carbonation oxygen-enriched leaching, wherein CO is obtained during leaching 2 The volume content of (2) is 85-87%, O 2 The volume content of (3) is 13-15%;
or leaching method 2: leaching vanadium in the calcified clinker by using an acid solution under the stirring condition to obtain a pickling solution containing vanadium and acid leaching waste slag vanadium, wherein the concentration of the acid and the time are as follows: the concentration of the acid solution is 35-65 wt.% within 0-10 min, and the concentration of the acid solution is less than 35wt.% and more than 10wt.% within 10-60 min;
the stirring is performed alternately 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, the intermediate product vanadium solution and CaCO are obtained by solid-liquid separation after leaching 3 And when the leaching method 2 is adopted for the main solid, the intermediate product vanadium solution and acid insoluble matters are obtained through solid-liquid separation after the leaching is finished.
The clockwise stirring and the anticlockwise stirring are alternately carried out, and a reaction tank with a rotatable base and a stirring system can be adopted, and the clockwise stirring and the anticlockwise rotation of the base are alternately carried out by the stirring system of the reaction tank. Or the counter-clockwise stirring and the clockwise rotation of the base of the reaction tank stirring system are alternately performed.
The roasting in the step a can be performed by heating with heat sources such as mixed gas, coke oven gas, electricity and the like.
The leaching reaction principle of the leaching method 1 mainly comprises the following steps:
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 for 1.5 to 3.5 hours at 800 to 850 ℃ to obtain roasted clinker; the step a comprises O 2 O in atmosphere of (2) 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 such that O 2 The volume ratio of the combustible gas to the air is 8-12: 10-20: 70-80.
In a specific embodiment, the combustible gas in the step a is coke oven gas or mixed gas; step a, caCO 3 Is limestone; the granularity of the vanadium slag is preferably 63-97 mu m, and the granularity of the limestone is preferably 125-200 mu m; the roasting in the step a is preferably roasting at 805-825 ℃; the roasting time is preferably 1.5-1.9 h; the CO adopted by the leaching in the step b 2 And O 2 Preferably C resulting from calcination in step aO 2 And O 2 。
In a specific embodiment, the leaching method 2 of step b: the concentration of the acid solution is 48-65 wt.% within 0-10 min, and the concentration of the acid solution is less than 35wt.% and more than 27wt.% within 10-60 min; preferably, the acid solution is added at a rate of: the relation between the sulfuric acid concentration C in the pickle liquor and the time t is maintained within 0-10 min:
maintaining the relation between the sulfuric acid concentration C and the time t in the pickle liquor for 10-60 min:
the acid solution is preferably added at a rate of 0.0015 to 0.085m 3 /min。
In a specific embodiment, the leaching method 2 of 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 8t; preferably, the ratio of the consumption of acid solution to the mass of calcified clinker during the leaching 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 anticlockwise stirring speed is 10-50 r/min; preferably, the clockwise stirring speed is 97r/min, and the anticlockwise stirring speed is 20r/min; after leaching by the leaching method 1 in the step b, the CaCO is used in the step c after primary acid leaching 3 The vanadium content in the main solid is preferably 0.6-1.05%; after leaching in leaching method 2, the content of vanadium in the acid insoluble matters in step c after primary acid leaching is 1.0-1.1 wt.%.
And c, adding ammonia water or ammonium salt into the vanadium solution for precipitating vanadium, wherein the principle of precipitating vanadium is as follows:
MnV 12 O 31 +2NH 4 + ＝(NH 4 ) 2 V 12 O 31 ↓+Mn 2+ 。
for example, the supernatant liquid without vanadium precipitation in the primary circulation is added with water, and the supernatant liquid with vanadium precipitation in the secondary circulation can be added with H 2 CO 3 And Mn of 2+ And the supernatant liquid of the vanadium precipitation can promote the carbonation oxygen-enriched leaching reaction.
In a specific embodiment, the leaching is a CO-containing process resulting from calcining step a 2 、O 2 Adding water or precipitating vanadium supernatant to leach out the tail gas of the roasting clinker, wherein the amount of the added water or precipitating vanadium supernatant is 1-5 times, preferably 1.5-2 times, of the amount of the roasting clinker; the supernatant fluid after vanadium precipitation of the vanadium solution in the step c is the supernatant fluid after vanadium precipitation, and the leaching time is preferably 60-120 min;
CaCO as described in step c 3 The main solid is returned to the step a for recycling.
The second purpose of the invention is to provide a green energy-saving method for extracting V, mn and Mg from calcined vanadium slag.
In order to achieve the second object of the invention, the method comprises the steps of calcining vanadium slag, immersing vanadium and carrying out solid-liquid separation to obtain insoluble matters and soluble matters;
the soluble substances are extracted by an extractant E 1 Extraction to obtain oil phase O 1 And aqueous phase A 1 The oil phase O 1 Back extraction phase separation is carried out by taking acid as a back extraction agent to obtain oil phase O 2 And aqueous phase A 2 Oil phase O 2 As extractant E 1 Recycling; the aqueous phase A 2 Vanadium is precipitated by adopting a vanadium precipitating agent, and then solid-liquid separation is carried out to obtain liquid L 1 And solids S 1 The liquid L 1 Can be used as back extraction agent for recycling, and solid S 1 Drying to obtain ammonium vanadate;
the aqueous phase A 1 By using extractant E 2 Extracting and phase-separating to obtain oil phase O 3 And aqueous phase A 3 Aqueous phase A 3 Recycling the vanadium leaching agent used as the vanadium leaching agent for the step a; oil phase O 3 Back extraction phase separation is carried out by adopting acid as a back extraction agent to obtain oil phase O 4 And aqueous phase A 4 Oil phase O 4 As extractant E 2 Recycling; aqueous phase A 4 Crystallizing above 60deg.C, and separating solid from liquid to obtain liquidBody L 2 And solids S 2 Solids S 2 Drying to obtain MnSO 4 ；
Liquid L 2 Recrystallizing below 40deg.C, and separating solid from liquid to obtain liquid L 3 And solids S 3 Liquid L 3 Can be used as back extraction agent for recycling, and solid S 3 Drying to obtain MgSO 4 ；
Wherein the extractant E 1 At least one of P204, P507, cyanex272, DNNSA, N1923, TOA, N263, TBP, naphthenic acid or tertiary carbonic acid, preferably P204; the extractant E 2 At least one of P204, P507, cyanex272, DNNSA, N1923, TOA, N263, TBP, naphthenic acid or tertiary carbonic acid, preferably P204;
the aqueous phase A 4 Preferably at above 100 ℃; liquid L 2 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 precipitating agent may be an existing conventional vanadium precipitating agent, for example, an ammonium salt, which may be (NH) 4 ) 2 SO 4 And NH 4 HCO 3 Etc.
The vanadium solution may be a vanadium solution obtained by other common calcination-vanadium leaching-solid-liquid separation methods, such as 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 includes V:0.6 to 1.8wt.%, mn:3.5 to 7.5wt.%, mg:0.3 to 0.6wt.%;
preferably, the vanadium slag is calcined and leached by the vanadium slag roasting leaching method, and the soluble matters are the vanadium solution in the step c.
In one embodiment, the extractant E 1 P204; the soluble substance adopts extractant E 1 The extraction ratio O/A of the extraction is preferably 1:1 to 1:10, more preferably 1:5;
aqueous phase A 1 By using extractant E 2 The extraction ratio O/A of the extraction is 2-3; preferably the aqueous phase A 1 By using 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 number of extraction is preferably 5 to 6;
the concentration of the stripping agent is preferably 20-60 wt.%;
the ratio O/A of the back extraction is preferably 10:1-1:1, more preferably 3:1;
the drying preferably utilizes the waste heat of the calcination in the step a.
The following describes the invention in more detail with reference to examples, which are not intended to limit the invention thereto.
Example 1
As shown in FIG. 1, vanadium slag is sieved by a 120-mesh sieve, 120-mesh undersize is sieved by a 160-mesh sieve, 160-mesh undersize (the granularity of 63-97 mu m is more than 95 percent and the granularity of 63 mu m is less than 5 percent) is evenly mixed with limestone powder of 80-mesh undersize, and then O is obtained 2 Roasting at 850 ℃ in an atmosphere with the content of 10 percent, the coke oven gas content of 15 percent and the air content of 75 percent to obtain roasted clinker for 2.5 hours; adding water into the baked clinker and calcining to generate CO 2 Performing carbonation oxygen-enriched leaching, wherein the water addition amount is 2.5 times of the roasting clinker amount, and controlling CO in the leaching atmosphere 2 Is 86% by volume, O 2 Is leached by stirring for 60min with the volume content of 14%; after leaching, the solid-liquid separation is carried out to obtain an intermediate product vanadium solution and CaCO 3 Mainly solid, caCO 3 The main solid returns to be used as limestone powder for recycling. CO 2 Leaching to realize zero corrosion of equipment sites.
The V concentration in the vanadium liquid reaches 36.31g/L, caCO is used 3 The main solid residue V was 0.95%.
Adding vanadium solution (NH) 4 ) 2 SO 4 And H 2 SO 4 And (3) vanadium precipitation, solid-liquid separation, and returning supernatant liquid of the vanadium precipitation to be used as leaching liquid.
Example 2
As shown in FIG. 1, the vanadium slag and the limestone powder are uniformly mixed and then are mixed in O 2 Roasting at 840 ℃ in an atmosphere with 12% of coke oven gas content and 70% of air content to obtain roasted clinker for 2.5 hours; CO produced by calcining the calcined clinker 2 Performing carbonation oxygen-enriched leaching water extraction, wherein the water addition amount is 2.5 times of the roasting clinker amount, and controlling CO in leaching atmosphere 2 Is 87% by volume, O 2 Is leached by stirring for 120min with the volume content of 15 percent; after leaching, the solid-liquid separation is carried out to obtain an intermediate product vanadium solution and CaCO 3 Mainly solid, caCO 3 The main solid returns to be used as limestone powder for recycling. CO 2 Leaching to realize zero corrosion of equipment sites. The V concentration in the vanadium liquid reaches 35.93g/L, caCO is used 3 The main solid residue V was 1.05%. Adding vanadium solution (NH) 4 ) 2 SO 4 And H 2 SO 4 And (3) vanadium precipitation, solid-liquid separation, and returning supernatant liquid of the vanadium precipitation to be used as leaching liquid.
Example 3
As shown in FIG. 1, the vanadium slag and the limestone powder are uniformly mixed and then are mixed in O 2 Roasting at 820 ℃ for 1.67 hours in an atmosphere with the content of 12 percent, the coke oven gas content of 15 percent and the air content of 73 percent to obtain roasted clinker; CO produced by calcining the calcined clinker 2 Performing carbonation oxygen-enriched leaching water extraction, wherein the water addition amount is 1.5 times of the roasting clinker amount, and controlling CO in leaching atmosphere 2 Is 87% by volume, O 2 Is leached by stirring for 80min with the volume content of 15 percent; after leaching, the solid-liquid separation is carried out to obtain an intermediate product vanadium solution and CaCO 3 Mainly solid, caCO 3 The main solid returns to be used as limestone powder for recycling. CO 2 Leaching to realize zero corrosion of equipment sites.
Roasting clinker to CO 2 After leaching, the V concentration in the vanadium liquid reaches 38.18g/L, caCO is used 3 The main solid residue V was 0.67%. Adding vanadium solution (NH) 4 ) 2 SO 4 And H 2 SO 4 And (3) vanadium precipitation, solid-liquid separation, and returning supernatant liquid of the vanadium precipitation to be used as leaching liquid.
Example 4
The preparation method of the roasting clinker is the same as that of the embodiment 1, leaching is carried out by adopting a device shown in fig. 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 calcified clinker is added, and the ratio is 0.075m 3 Adding 55% sulfuric acid dropwise at a flow rate of/min for 10min, and adding 0.0015m 3 Sulfuric acid with the concentration of 10% is added dropwise at the flow rate of/min. And stopping acid leaching after acid leaching for 60 min. The vanadium content in the acid leaching residue was 1.05%.
The acid consumption is 0.825m 3 The pot is not sticky during the reaction.
Example 5
The preparation method of the roasting clinker is the same as that of the embodiment 1, leaching is carried out by adopting a device shown in fig. 3, clockwise stirring and anticlockwise stirring are alternately carried out, the clockwise stirring speed is 97r/min, the anticlockwise stirring speed is 20r/min, 7.0t calcified clinker is added, and the leaching speed is 0.085m 3 Adding 65% sulfuric acid dropwise at a flow rate of/min for 10min, and adding 0.0025m 3 Sulfuric acid with the concentration of 27% is added dropwise at the flow rate of/min. And stopping acid leaching after acid leaching for 60 min. The vanadium content in the acid leaching residue was 1.08%.
The acid consumption is 0.975m 3 The pot is not sticky during the reaction.
Example 6
The technological process of the embodiment is shown in figure 8, a, roasting vanadium slag for 4 hours at 900 ℃ by adopting the conventional roasting acid leaching, leaching for 60 minutes at pH 2.8-3.2, performing calcification calcination on the vanadium slag, performing vanadium leaching and solid-liquid separation by adopting dilute sulfuric acid with the mass concentration of 15% to obtain insoluble matters and soluble matters, and measuring V25.40 g/L, mn 11.10g/L and Mg 1.68g/L in the soluble matters; 1.52wt.% V, 4.11wt.% Mn, 0.58wt.% Mg in insoluble matter;
b. the soluble substances are extracted and split by P204 to obtain oil phase O 1 And aqueous phase A 1 The ratio of O/A is 1:5, the times of extraction are 5 times, the extraction rate of vanadium is 100%, the extraction rate of manganese is 89.84%, and the extraction rate of magnesium is 81.54%. Oil phase O 1 The dilute sulfuric acid with the mass concentration of 20 percent is used as a stripping agent to carry out stripping phase separation to obtain oil phase O 2 And aqueous phase A 2 Oil phase O 2 Recycling as extractant P204; aqueous phaseA 2 By means of (NH) 4 ) 2 SO 4 And H 2 SO 4 Vanadium precipitation and solid-liquid separation are carried out to obtain liquid L 1 And solids S 1 Liquid L 1 Can be used as back extraction agent for recycling, and solid S 1 Drying the red cake by using the calcined waste heat to obtain ammonium vanadate;
c. aqueous phase A 1 Extracting and phase-separating by P15 for 5 times to obtain oil phase O 3 And aqueous phase A 3 The ratio of O/A is 3:1, the extraction rate of Mn and Mg is shown in Table 1, and the water phase A is 3 Recycling as a vanadium leaching agent in the step a; oil phase O 3 Back extraction phase separation is carried out by adopting dilute sulfuric acid with the mass concentration of 20 percent to obtain oil phase O 4 And aqueous phase A 4 Back extracted O/A3:1, oil phase O 4 The water is used as an extractant P15 for recycling; aqueous phase A 4 Crystallizing at 100deg.C, and separating solid from liquid to obtain liquid L 2 And solids S 2 Solids S 2 Drying by using calcined waste heat to obtain MnSO 4 ；
d. Liquid L 2 Crystallizing at 25deg.C, and separating solid from liquid to obtain liquid L 3 And solids S 3 Liquid L 3 Can be used as back extraction agent for recycling, and solid S 3 Drying by using calcined waste heat to obtain MgSO 4 。
Table 1 extraction yield of Mn and Mg by five-stage extraction of P15
Extraction stage number 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 example 1, the green energy-saving method for efficiently extracting V, mn and Mg by adopting the solid waste reduction grading without waste liquid discharge of the calcined vanadium slag of the invention can not only obtain ammonium vanadate and MnSO 4 ，MgSO 4 Three qualified products, and can also make the extractant P204 and P15 and the acid leaching agent and the back extractant H 2 SO 4 、HNO 3 、H 3 PO 4 The waste heat of the roasting, HCl and the like are recycled, 6 large cycles of zero discharge of an extractant, zero discharge of a vanadium leaching agent, zero discharge of a back extraction agent, zero discharge of waste heat and the like are formed, zero generation of gypsum slag is realized, no waste liquid discharge-solid waste reduction-waste heat utilization is realized, thereby improving the utilization rate of vanadium, manganese and magnesium resources, 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 by about 40 percent from zero, and the Mg resource utilization rate is improved by about 20 percent from zero.
Comparative example 1
As shown in FIG. 2, the vanadium slag and the limestone powder are uniformly mixed and then mixed in O 2 Roasting at 840 ℃ in an atmosphere with 12% of coke oven gas content, 18% of air content and 70% of air content to obtain roastingClinker for 2.5 hours; leaching the roasted clinker with dilute sulfuric acid with 65% concentration, 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 120min; solid-liquid separation after leaching is completed to obtain intermediate product vanadium solution and CaSO 4 Mainly solid waste. The V concentration in the vanadium liquid reaches 27.58g/L, caCO is used 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 process, the acid consumption was 1.2m 3 . The vanadium content in the acid leaching residue was 1.53wt.%. The reaction process is shown in FIG. 6 as can adhesive.
Comparative example 3
Otherwise, the same as in example 4, except that dilute sulfuric acid having a concentration of 10% was used throughout the process, the acid consumption was 1.5m 3 . The vanadium content in the acid leaching residue was 1.69wt.%. The reaction process is shown in FIG. 6 as can adhesive.
Comparative example 4
Otherwise, the same as in example 4 was conducted except that the stirring was conducted by stirring clockwise at a stirring speed of 90r/min and an acid consumption of 1.3m 3 . The vanadium content in the acid leaching residue was 1.71wt.%. The fine material is stuck on the bottom without running away.

Claims (31)

1. The vanadium slag roasting and leaching method is characterized by comprising the following steps of:
a. mixing vanadium slag with CaCO 3 Mixing uniformly and then adding O 2 Roasting in the atmosphere of (2) to obtain roasting clinker; the step a comprises O 2 O in atmosphere of (2) 2 The volume content of the gas is 8-32%, and the rest gas in the atmosphere is at least one of nitrogen, inert gas and combustible gas;
b. leaching method 1: CO for the roasted clinker 2 Performing carbonation oxygen-enriched leaching, wherein CO is obtained during leaching 2 The volume content of (2) is 85-87%, O 2 The volume content of (3) is 13-15%; the leaching is that CO is generated by calcining the step a 2 、O 2 Adding water or precipitating vanadium supernatant to leach the tail gas and the roasting clinker, wherein the amount of the added water or precipitating vanadium supernatant is 1 to the range of the amount of the roasting clinkerC, precipitating vanadium in the vanadium solution, wherein the vanadium precipitation supernatant is 5 times that of the vanadium solution obtained after the vanadium precipitation in the step c;
or leaching method 2: leaching vanadium in the calcified clinker by using an acid solution under the stirring condition to obtain a pickling solution containing vanadium and acid leaching waste slag vanadium, wherein the concentration of the acid and the time are as follows: the concentration of the acid solution is 35-65 wt.% within 0-10 min, and the concentration of the acid solution is less than 35wt.% and more than 10wt.% within 10-60 min; the acid solution is added at the following rate: the relation between the sulfuric acid concentration C in the pickle liquor and the time t is maintained within 0-10 min:
maintaining the relation between the sulfuric acid concentration C and the time t in the pickle liquor for 10-60 min:
the stirring is performed alternately by clockwise stirring and anticlockwise stirring; the acid solution is at least one of sulfuric acid, hydrochloric acid, nitric acid or phosphoric acid; b, the clockwise stirring speed is 85-100 r/min, and the anticlockwise stirring speed is 10-50 r/min;
c. when the leaching method 1 is adopted, the intermediate product vanadium solution and CaCO are obtained by solid-liquid separation after leaching 3 And when the leaching method 2 is adopted for the main solid, the intermediate product vanadium solution and acid insoluble matters are obtained through solid-liquid separation after the leaching is finished.
2. The vanadium slag roasting and leaching method according to claim 1, wherein the roasting in the step a is carried out for 1.5-3.5 hours at 800-850 ℃ to obtain roasting clinker; the atmosphere in the step a is O 2 The volume ratio of the combustible gas to the air is 8-12: 10-20: 70-80.
3. The vanadium slag roasting and leaching method according to claim 1 or 2, characterized in thatThe combustible gas in the step a is coke oven gas and mixed gas; step a, caCO 3 Is limestone.
4. The vanadium slag roasting and leaching method according to claim 3, wherein the granularity of the vanadium slag is 63-97 μm.
5. A vanadium slag roasting and leaching method according to claim 3, wherein the limestone has a particle size of 125-200 μm.
6. The method according to claim 3, wherein the roasting in the step a is a roasting at 805-825 ℃.
7. The method for roasting and leaching vanadium slag according to claim 3, wherein the roasting time in the step a is 1.5-1.9 h.
8. The vanadium slag roasting and leaching method according to claim 1 or 2, characterized by the leaching method 2 of step b: the concentration of the acid solution is 48-65 wt.% within 0-10 min, and the concentration of the acid solution is less than 35wt.% and more than 27wt.% within 10-60 min.
9. The vanadium slag roasting and leaching method according to claim 8, wherein the acid solution is added at a rate of 0.0015 to 0.085m 3 /min。
10. The vanadium slag roasting and leaching method according to claim 1 or 2, characterized by the leaching method 2 of 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 /8t。
11. The vanadium slag roasting and leaching method according to claim 10, wherein the ratio of the consumption of acid solution to the mass of calcified clinker in the leaching process is 0.8m 3 /6.5t～1.0m 3 /7t。
12. The vanadium slag roasting and leaching method according to claim 1 or 2, wherein the clockwise stirring speed in the step b is 97r/min, and the anticlockwise stirring speed is 20r/min.
13. The method according to claim 12, wherein after leaching in step b, the leaching method 1 is adopted, the step c is adopted after primary acid leaching 3 The vanadium content in the main solid is 0.6-1.05%; after leaching in leaching method 2, the content of vanadium in the acid insoluble matters in step c after primary acid leaching is 1.0-1.1 wt.%.
14. The vanadium slag roasting and leaching method according to claim 1 or 2, wherein the amount of the added water or the precipitated vanadium supernatant is 1.5-2 times of the amount of the roasting clinker;
CaCO as described in step c 3 The main solid is returned to the step a for recycling.
15. The vanadium slag roasting and leaching method according to claim 14, wherein the leaching time is 60 to 120min.
16. The method for extracting V, mn and Mg from the calcined vanadium slag in a green and energy-saving way is characterized by comprising the steps of calcining the vanadium slag, immersing vanadium and carrying out solid-liquid separation to obtain insoluble matters and soluble matters; the vanadium slag is calcined and leached by the vanadium slag roasting and leaching method according to any one of claims 1 to 15, and the soluble matters are the vanadium solution in the step c;
the soluble substances are extracted by an extractant E 1 Extraction to obtain oil phase O 1 And aqueous phase A 1 The oil phase O 1 Back extraction phase separation is carried out by taking acid as a back extraction agent to obtain oil phase O 2 And aqueous phase A 2 Oil phase O 2 As extractant E 1 Recycling; the aqueous phase A 2 Vanadium is precipitated by adopting a vanadium precipitating agent, and then solid-liquid separation is carried out to obtain liquid L 1 And solids S 1 The liquidL 1 Can be used as back extraction agent for recycling, and solid S 1 Drying to obtain ammonium vanadate;
the aqueous phase A 1 By using extractant E 2 Extracting and phase-separating to obtain oil phase O 3 And aqueous phase A 3 Aqueous phase A 3 Recycling the vanadium leaching agent used as the vanadium leaching agent for the step a; oil phase O 3 Back extraction phase separation is carried out by adopting acid as a back extraction agent to obtain oil phase O 4 And aqueous phase A 4 Oil phase O 4 As extractant E 2 Recycling; aqueous phase A 4 Crystallizing at above 60deg.C, and separating solid from liquid to obtain liquid L 2 And solids S 2 Solids S 2 Drying to obtain MnSO 4 ；
Liquid L 2 Recrystallizing below 40deg.C, and separating solid from liquid to obtain liquid L 3 And solids S 3 Liquid L 3 Can be used as back extraction agent for recycling, and solid S 3 Drying to obtain MgSO 4 ；
Wherein the extractant E 1 At least one of P204, P507, cyanex272, DNNSA, N1923, TOA, N263, TBP, naphthenic acid or tertiary carbonic acid; the extractant E 2 Is at least one of P204, P507, cyanex272, DNNSA, N1923, TOA, N263, TBP, naphthenic acid or tertiary carbonic acid.
17. The method for extracting V, mn and Mg from calcined vanadium slag according to claim 16, wherein the extractant E 1 P204.
18. The method for extracting V, mn and Mg from calcined vanadium slag according to claim 16, wherein the extractant E 2 P204.
19. The method for extracting V, mn and Mg from calcined vanadium slag according to claim 16, wherein the water phase A 4 Crystallizing at above 100deg.C.
20. The green energy saving extraction of V, mn from calcined vanadium slag as claimed in claim 16A method for Mg, characterized in that the liquid L 2 Recrystallizing at below 25deg.C.
21. The method for extracting V, mn and Mg from calcined vanadium slag according to claim 16, wherein the stripping agent is H 2 SO 4 、HNO 3 、H 3 PO 4 Or HCl.
22. The method for extracting V, mn and Mg from the calcined vanadium slag according to claim 21, wherein the stripping agent is sulfuric acid.
23. The method for extracting V, mn, mg from calcined vanadium slag according to claim 16, wherein the soluble materials include V: 15-45 g/L, mn: 10-20 g/L, mg: 1-5 g/L; the acid insoluble matter includes V:0.6 to 1.8wt.%, mn:3.5 to 7.5wt.%, mg:0.3 to 0.6wt.%.
24. The method for extracting V, mn and Mg from calcined vanadium slag according to claim 16, wherein the extractant E 1 P204; the soluble substance adopts extractant E 1 The extraction ratio O/A of the extraction is 1:1-1:10;
aqueous phase A 1 By using extractant E 2 The ratio of O/A in the extraction is 2-3.
25. The method for extracting V, mn and Mg from calcined vanadium slag according to claim 24, wherein the soluble substances adopt an extractant E 1 The extraction ratio O/A of the extraction is 1:5.
26. The method for extracting V, mn, mg from calcined vanadium slag according to claim 24, wherein the aqueous phase a 1 By using 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 is42.69% -81.54%.
27. The method for extracting V, mn, mg from vanadium slag by green energy saving according to claim 24, wherein the number of the extractions is 5 to 6.
28. The method for extracting V, mn, mg from calcined vanadium slag according to claim 24, wherein the concentration of the stripping agent is 20% to 60% by weight.
29. The method for extracting V, mn and Mg from the calcined vanadium slag according to claim 24, wherein the counter extraction ratio O/A is 10:1-1:1.
30. The method for extracting V, mn, mg from vanadium slag by green energy saving according to claim 29, wherein the ratio O/a of the stripping is 3:1.
31. The method for extracting V, mn, and Mg from vanadium slag by green energy saving according to claim 24, wherein said drying utilizes the waste heat of the calcination in step a.
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