CN115198116A - Method for extracting vanadium from vanadium-containing stone coal - Google Patents
Method for extracting vanadium from vanadium-containing stone coal Download PDFInfo
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- CN115198116A CN115198116A CN202210863634.4A CN202210863634A CN115198116A CN 115198116 A CN115198116 A CN 115198116A CN 202210863634 A CN202210863634 A CN 202210863634A CN 115198116 A CN115198116 A CN 115198116A
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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
- C22B34/00—Obtaining refractory metals
- C22B34/20—Obtaining niobium, tantalum or vanadium
- C22B34/22—Obtaining vanadium
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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
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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
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
- C22B3/08—Sulfuric acid, other sulfurated acids or salts thereof
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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
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
- C22B3/42—Treatment or purification of solutions, e.g. obtained by leaching by ion-exchange extraction
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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
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
- C22B3/44—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes
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- 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
Abstract
The invention discloses a method for extracting vanadium from vanadium-containing stone coal, which comprises the following steps: (1) roasting and leaching: roasting and leaching the vanadium-containing stone coal powder to obtain vanadium-containing leaching solution; (2) enriching: enriching vanadium in the vanadium-containing leaching solution to obtain a vanadium-enriched solution; (3) aging: adjusting the pH value of the vanadium-enriched liquid to 8.5-9.5, standing, aging and filtering to obtain an aged filtrate; (4) removing impurities: and adding an impurity removing agent into the aged filtrate, and carrying out heat preservation, standing and filtering to obtain the vanadium-containing purified liquid, wherein the impurity removing agent comprises a chelating agent and a flocculating agent. The method has the advantages of simple process control, stable product quality, shortened process flow, reduced ammonia nitrogen wastewater generation amount from the source, saved raw materials, reduced energy consumption, controlled vanadium loss, accordance with the requirement of clean production, and suitability for large-scale production.
Description
Technical Field
The invention belongs to the technical field of vanadium extraction and purification, and relates to a process for directly and deeply removing impurities and extracting vanadium by using vanadium-containing stone coal as a raw material through vanadium enrichment liquid.
Background
The stone coal is one of the main raw materials for extracting vanadium pentoxide as a vanadium-containing polymetallic paragenetic mineral. In the existing process for preparing high-purity vanadium pentoxide from vanadium-containing stone coal, a vanadium-enriched liquid of the vanadium-containing stone coal is usually subjected to impurity removal by a flocculating agent, ammonium chloride is used for precipitating vanadium to obtain a crude product of ammonium metavanadate, an impurity removing agent is used for secondary deep impurity removal after the ammonium metavanadate is subjected to alkali dissolution, and ammonium chloride is added for precipitating vanadium to obtain high-purity ammonium metavanadate. The methods have the defects of long flow, high vanadium loss, large amount of ammonia nitrogen wastewater generated by secondary vanadium precipitation of ammonium chloride, about 40 tons of ammonia nitrogen wastewater generated by preparation of each ton of high-purity vanadium pentoxide, large treatment capacity of the ammonia nitrogen wastewater, difficulty in comprehensive utilization of salt slag generated by conversion of the added ammonium chloride, high production and operation cost, large medicament consumption, low wastewater circulation rate, large environment-friendly treatment pressure and the like.
Disclosure of Invention
In order to overcome the defects of the prior art, the invention aims to provide a process for directly and deeply removing impurities and extracting vanadium from a vanadium-enriched liquid, wherein aging, a flocculating agent and a chelating agent are used for deeply removing impurities, the purified liquid after the impurities are removed can obtain high-purity ammonium polyvanadate in an ammonia-free vanadium precipitation mode, and a vanadium precipitation tail water solution can be recycled to a production system. The method does not need the lengthy flow of secondary impurity removal and secondary ammonium chloride vanadium precipitation, has relatively low energy consumption and small vanadium loss, reduces the consumption of raw materials and obtains the qualified high-purity vanadium pentoxide product.
Specifically, the invention provides a method for extracting vanadium from vanadium-containing stone coal, which comprises the following steps:
(1) Roasting and leaching: roasting and leaching the vanadium-containing stone coal powder to obtain vanadium-containing leaching solution;
(2) Enrichment: enriching vanadium in the vanadium-containing leaching solution to obtain a vanadium-enriched solution;
(3) Aging: adjusting the pH value of the vanadium-enriched liquid to 8.5-9.5, standing, aging and filtering to obtain an aged filtrate;
(4) Removing impurities: and adding an impurity removing agent into the aged filtrate, and carrying out heat preservation, standing and filtering to obtain the vanadium-containing purified liquid, wherein the impurity removing agent comprises a chelating agent and a flocculating agent.
In one or more embodiments, in the step (1), the roasting includes performing sulfur-fixing activation roasting, and then adding a calcium compound and a barium compound to the roasted slag after the sulfur-fixing activation roasting, and then performing roasting.
In one or more embodiments, in step (1), the leaching comprises soaking the roasted mass with an acidic solution, preferably a sulfuric acid solution.
In one or more embodiments, in step (2), the enrichment is performed by adsorption on an anion exchange resin, preferably a D301 resin.
In one or more embodiments, in the step (2), the vanadium concentration of the vanadium-enriched liquid is greater than or equal to 60gV 2 O 5 /L。
In one or more embodiments, in step (2), the anion exchange resin having adsorbed vanadium is desorbed using a caustic solution.
In one or more embodiments, in step (3), the pH of the vanadium-enriched liquor is adjusted to 8.5-9.5 using sulfuric acid.
In one or more embodiments, in step (3), the standing aging time is not less than 48h.
In one or more embodiments, in the step (4), the temperature is kept at 60-80 ℃ after the impurity removing agent is added, and the standing time is 4-6h.
In one or more embodiments, in step (4), the chelating agent is ethylenediaminetetraacetic acid.
In one or more embodiments, in step (4), the flocculant is one or more selected from the group consisting of aluminum sulfate, magnesium chloride, and aluminum chloride.
In one or more embodiments, in step (4), the chelating agent is used in an amount of 0.1% to 1% by mass of the filtrate.
In one or more embodiments, in step (4), the amount of flocculant used is 1% to 10% of the mass of the filtrate.
In one or more embodiments, in step (4), the chelating agent is added, and the mixture is stirred uniformly, for example, for 10-30min, and then the flocculating agent is added.
In one or more embodiments, the method further comprises the steps of:
(5) Precipitating vanadium: precipitating vanadium in the vanadium-containing purification solution in the form of ammonium polyvanadate, and carrying out solid-liquid separation to obtain the ammonium polyvanadate.
In one or more embodiments, in the step (5), the pH value of the vanadium-containing purification solution is adjusted to 1.5-2.1, the vanadium-containing purification solution is heated to 80-100 ℃, and the vanadium is precipitated in the form of ammonium polyvanadate by stirring.
In one or more embodiments, the method further comprises the steps of:
(6) And (3) calcining: calcining the ammonium polyvanadate to decompose the ammonium polyvanadate to obtain vanadium pentoxide.
In one or more embodiments, in step (6), the calcination temperature is from 500 to 550 ℃ and the calcination time is from 3 to 5 hours.
Detailed Description
To make the features and effects of the present invention comprehensible to those skilled in the art, general description and definitions are made below with reference to terms and expressions mentioned in the specification and claims. Unless defined otherwise, all technical and scientific terms used herein have the same meaning as commonly understood by one of ordinary skill in the art to which this invention belongs.
The theory or mechanism described and disclosed herein, whether correct or incorrect, should not limit the scope of the present invention in any way, i.e., the present disclosure may be practiced without being limited by any particular theory or mechanism.
The terms "comprising," "including," "containing," and the like, herein, encompass the meanings of "consisting essentially of 8230 \8230%, \8230composition" and "consisting of 8230 \823030composition," for example, when "a comprises B and C" is disclosed herein, "a consists essentially of B and C" and "a consists of B and C" should be considered to have been disclosed herein.
All features defined herein as numerical ranges or percentage ranges, such as values, amounts, levels and concentrations, are for brevity and convenience only. Accordingly, the description of numerical ranges or percentage ranges should be considered to cover and specifically disclose all possible subranges and individual numerical values (including integers and fractions) within the range.
Herein, unless otherwise specified, percentages refer to mass percentages and ratios to mass ratios.
Herein, when embodiments or examples are described, it is to be understood that they are not intended to limit the invention to these embodiments or examples. On the contrary, all alternatives, modifications, and equivalents of the methods and materials described herein are intended to be included within the scope of the invention as defined by the appended claims.
In the present context, for the sake of brevity, all possible combinations of various features in various embodiments or examples are not described. Therefore, the respective features in the respective embodiments or examples may be arbitrarily combined as long as there is no contradiction between the combinations of the features, and all the possible combinations should be considered as the scope of the present specification.
The method for extracting vanadium from vanadium-containing stone coal comprises the following steps: roasting, leaching, enriching, aging and removing impurities.
Before roasting and leaching, the vanadium-containing stone coal raw ore can be pretreated to obtain vanadium-containing stone coal powder. The pre-processing may include: crushing the stone coal raw ore, and screening to obtain vanadium-containing stone coal powder. The calorific value of the stone coal raw ore can be 800-1200 kcal/kg. V of stone coal raw ore 2 O 5 The content may be 0.85wt% to 0.9wt%, for example 0.87wt%. The crushing may be two-stage crushing, such as coarse crushing in a jaw crusher and crushing in a combined crusher. Can be used forSieving with a vibrating sieve. The preferred particle diameter of the vanadium-containing stone coal powder applicable to the invention is less than or equal to 8mm.
In the present invention, preferably, the roasting includes first performing sulfur fixation activation roasting, and then adding a calcium compound and a barium compound to the roasted slag after the sulfur fixation activation roasting, and then performing roasting.
In the sulfur-fixing activation roasting, caO is added into the vanadium-containing stone coal powder according to the sulfur content, and then roasting is carried out. The mass ratio of the addition amount of CaO to the sulfur content of the pulverized stone coal can be 2. The time for activating and roasting the sulfur can be 40-100min, preferably 40-80min, for example about 60 min. The temperature of the sulfur-fixing activation roasting can be 800-1000 ℃, preferably 850-950 ℃, for example, about 900 ℃. The sulfur fixation activation roasting may be carried out in a boiler, preferably a circulating fluidized bed boiler. Preferably, the oxygen content in the flue gas at the outlet of the boiler is controlled to be more than or equal to 5vol% in the sulfur-fixing activating roasting.
Collecting the cinder (including bottom cinder and cinder) after sulfur-fixing activation roasting, adding calcium compound and barium compound, and then carrying out activation roasting to obtain the roasted clinker. The activation firing is preferably carried out in a tunnel kiln. Preferably, after the calcium compound and the barium compound are added, the mixture is crushed into 60-100 meshes, preferably about 80 meshes, and then the brick is made and the blank is stacked, and then the mixture enters a tunnel kiln for high-temperature oxygen-enriched roasting. The firing temperature (e.g., the temperature in the high temperature zone of the tunnel kiln) may be 800 to 1000 deg.C, preferably 850 to 950 deg.C, e.g., about 900 deg.C. The calcination time (e.g., high temperature residence time in a tunnel kiln) is 3 to 5 hours, preferably 3.5 to 4.5 hours, e.g., about 4 hours. The total residence time in the tunnel kiln may be in the range of from 6 to 12 hours, preferably from 8 to 10 hours, for example around 9 hours. The oxygen content in the flue gas generated by roasting is preferably more than or equal to 10vol%.
The calcium compound suitable for use in the present invention may be an oxide of calcium, preferably CaO. The amount of the calcium compound added is preferably 1% to 5%, for example, 2%, 3%, or 4% of the mass of the clinker after the sulfur-fixing activation roasting. The barium compound suitable for use in the present invention may be a barium salt, preferably BaSO 4 . The amount of the barium compound added is preferably 2 to 10% by mass, for example, 3%, 4%, 5%, 6%, 8% by mass of the roasted slag after sulfur-fixing activation roasting。
In the leaching, the roasted clinker obtained by roasting is stirred and leached to obtain vanadium-containing leaching solution. Before leaching, the roasted clinker may be first crushed to a particle size of 40-80 mesh, for example 50 mesh, 60 mesh, 70 mesh. The roasted clinker may be soaked with an acidic solution (e.g. an aqueous acid solution), preferably a sulphuric acid solution, as a leaching agent. The liquor ratio (mass ratio of roasting clinker to leaching agent) of the soaking system can be 1. The pH value of the soaking system is preferably 1.4-2.5. In some embodiments, a sulfuric acid solution is used as the leaching agent, the addition amount of sulfuric acid is 3% to 5%, for example 4%, of the mass of the roasted clinker, and the feed-to-liquor ratio is 1. The leaching time can be 30min-90min, such as 45min, 60min, 75min.
According to the invention, the roasting is carried out by firstly carrying out sulfur fixing activation roasting and then adding the calcium compound and the barium compound, so that the total leaching rate of vanadium in the stone coal can be improved, and the total leaching rate of vanadium is far higher than that of a conventional roasting vanadium extraction method. In some embodiments, the total leaching rate of vanadium in the stone coal is 85% or more, such as 87% or more, 87.8%.
And after the leaching is finished, carrying out solid-liquid separation on the leached feed liquid to obtain vanadium-containing leaching liquid and slag. The solid-liquid separation can be carried out by adopting a mode of firstly filtering by using a belt type vacuum filter and then carrying out fine pressing on the filtrate by using a box type filter press. The slag may be washed with water, and for example, the slag may be washed with ion exchange tail water produced in the subsequent ion exchange adsorption. The wash water from the washing can be used in the leaching process.
In the enrichment step, vanadium in the vanadium-containing leaching solution is enriched to obtain an enriched vanadium solution. The enrichment can be carried out by means of ion exchange adsorption. In some embodiments, the enrichment is performed by adsorption on an anion exchange resin. The anion exchange resin is preferably a D301 resin. In the invention, the classification and naming rules of the ion exchange resin are in accordance with the GB/T1631-2008 ion exchange resin naming system and basic specifications. Preferably, the anion exchange resin adsorbed with vanadium is desorbed using a lye. The vanadium concentration of the vanadium-enriched liquid is excellentSelecting more than or equal to 60gV 2 O 5 L,. Gtoreq.80 gV 2 O 5 /L、81.2g V 2 O 5 /L、82.6gV 2 O 5 /L、85.3gV 2 O 5 And L. The alkali in the alkaline solution can be selected from Na 2 CO 3 、NaHCO 3 And NaOH, preferably NaOH. The concentration of the lye may be from 50g/L to 100g/L, preferably from 70g/L to 90g/L, for example 80g/L. The vanadium concentration of the liquid to be eluted reaches the process requirement, for example, more than or equal to 60gV 2 O 5 L, the desorption process can be stopped.
In the aging step, the pH value of the vanadium-enriched liquid is adjusted to 8.5-9.5, such as about 9, and the aged liquid is obtained after standing, aging and filtering. The pH can be adjusted by adding an acidic substance, preferably sulfuric acid, to the vanadium-enriched liquor. The standing and aging time is usually more than or equal to 48 hours, such as 60 hours, 72 hours and 96 hours. The purpose of aging is to precipitate out the silicon dioxide in the vanadium-enriched liquid.
In the impurity removal step, an impurity removal agent is added into the aged filtrate, and the vanadium-containing purified solution is obtained through heat preservation, standing and filtration. In the invention, the impurity removing agent comprises a chelating agent and a flocculating agent.
The chelating agent suitable for use in the present invention is preferably ethylenediaminetetraacetic acid (EDTA). The amount of chelating agent used is preferably 0.1% to 1%, for example 0.2%, 0.3%, 0.4%, 0.5%, 0.6%, 0.7%, 0.8%, 0.9% by mass of the filtrate after aging. The flocculating agent suitable for the invention is one or more selected from aluminum sulfate, magnesium chloride and aluminum chloride. The amount of flocculating agent is preferably 1% to 10%, e.g. 2%, 3%, 4%, 5%, 6%, 7%, 8% of the mass of the filtrate after ageing. In some embodiments, the chelating agent is ethylenediaminetetraacetic acid, the chelating agent is present in an amount of 0.4% to 0.6%, e.g., 0.5%, by mass of the filtrate after aging, the flocculating agent is magnesium chloride, and the flocculating agent is present in an amount of 3% to 5%, e.g., 4%, by mass of the filtrate after aging. In some embodiments, the chelating agent is ethylenediaminetetraacetic acid, the chelating agent is present in an amount of 0.1% to 0.3%, e.g., 0.2%, by mass of the aged filtrate, the flocculating agent is aluminum sulfate and aluminum chloride, the aluminum sulfate is present in an amount of 3% to 5%, e.g., 4%, by mass of the aged filtrate, and the aluminum chloride is present in an amount of 1% to 3%, e.g., 2%, by mass of the aged filtrate. In some embodiments, the chelating agent is ethylenediaminetetraacetic acid, the chelating agent is present in an amount of 0.5% to 0.7%, e.g., 0.6%, by mass of the filtrate after aging, the flocculating agent is aluminum sulfate, and the flocculating agent is present in an amount of 5% to 7%, e.g., 6%, by mass of the filtrate after aging.
In a preferred embodiment, the chelating agent is added, the mixture is stirred uniformly, then the flocculating agent is added, the mixture is stirred uniformly, and then the mixture is kept warm and kept stand. Preferably, the chelating agent is added and then stirred for 10-30min, such as 15min, 20min, 25min. The addition mode of the chelating agent and the flocculating agent is favorable for improving the impurity removal effect.
After the addition of the impurity-removing agent, the temperature of the mixture is preferably 60 to 80 ℃ such as 65 ℃,70 ℃ or 75 ℃. The time for the incubation and standing is preferably 4 to 6 hours, for example, 4.5 hours, 5 hours, 5.5 hours. After standing at the temperature, the solution may be cooled, for example to about 40 ℃, and then filtered. Filtration can be carried out using a filter press. The filtrate (i.e. vanadium-containing purified liquid) obtained by filtering can be used for vanadium precipitation reaction.
The method for extracting vanadium from vanadium-containing stone coal can also comprise a vanadium precipitation step.
In the vanadium precipitation step, vanadium in the vanadium-containing purification solution obtained after impurity removal is precipitated in the form of ammonium polyvanadate, and the ammonium polyvanadate is obtained through solid-liquid separation. In some embodiments, the pH of the vanadium-containing purification solution is adjusted to 1.5-2.1, e.g., 1.6, 1.8, 2.0, heated to 80-100 deg.C, e.g., 90 deg.C, and stirred to precipitate the vanadium in the liquid as ammonium polyvanadate. The vanadium precipitation can be carried out in a vanadium precipitation reaction kettle. The solid-liquid separation can be carried out by adopting a centrifugal mode. The vanadium precipitation tail water obtained by solid-liquid separation can be reused in the leaching process.
The method for extracting vanadium from vanadium-containing stone coal can also comprise a calcination step.
In the calcining step, ammonium polyvanadate is heated and decomposed to obtain vanadium pentoxide. The calcination temperature may be 500-550 ℃. The calcination time may be 3-5h, e.g. 4h. The calcination may be carried out in a rotary kiln. The method can be used for preparing high-purity vanadium pentoxide. In the invention, the high-purity vanadium pentoxide means that the content of the vanadium pentoxide is more than 99.5 wt%.
The invention has the following beneficial effects:
the method has the advantages of simple process control, stable product quality, shortened process flow, reduced ammonia nitrogen wastewater generation amount from the source, raw material saving, energy consumption reduction, vanadium loss control, accordance with the requirement of clean production, and suitability for large-scale production. Through detection, the purity of the high-purity vanadium pentoxide prepared by the process can reach over 99.5 percent, the impurity content can be less than 100ppm, and the high-purity vanadium pentoxide can be used as the special high-purity vanadium pentoxide for the vanadium electrolyte.
The present invention will be further described below by way of specific examples. It should be understood that these examples are illustrative only and are not intended to limit the scope of the present invention. The methods and reagents used in the examples are, unless otherwise indicated, conventional in the art.
Example 1
1. Decarburization in a fluidized bed furnace: containing V 2 O 5 0.87wt% of stone coal raw ore (calorific value of 800-1200 Kcal/kg) is coarsely crushed by a jaw crusher, then enters a combined crusher for crushing, is screened by a vibrating screen, the ore particles are controlled to be less than 8mm, and unqualified materials are returned by a belt conveyor. Adding lime with the sulfur content of 3 times of that of the ore into the crushed materials, detecting that the sulfur content of the ore is 0.68wt%, adding CaO with the sulfur content of 2wt%, uniformly mixing, feeding the mixture into a circulating fluidized bed boiler by a feeder to perform sulfur fixation, activation and roasting, wherein the roasting time is about 60min, the roasting temperature in the boiler is 900 ℃, and controlling the oxygen content in the smoke at the outlet of the boiler: more than or equal to 5vol%, and the boiler flue gas is discharged after being treated in a centralized manner.
2. Activating and roasting in a tunnel kiln: collecting the bottom slag and ash slag after sulfur fixation activation roasting in the circulating fluidized bed boiler, adding CaO accounting for 2 percent of the mass of the low slag and ash slag and BaSO accounting for 4 percent of the mass of the low slag and ash slag 4 Crushing the mixture to about 80 meshes by using a ball mill, making the brick by using a double-machine vacuum extruder, stacking the brick to a kiln car, and entering a tunnel kiln for high-temperature oxygen-enriched roasting. Residence time in the high temperature zone (temperature 900 ℃): 4 hours, the total retention time is 9 hours, and the oxygen content in the flue gas is as follows: more than or equal to 10vol%, and the flue gas is discharged after reaching standards after being treated in a centralized way.
3. Continuous leaching of roasted clinker: crushing the roasted clinker to 60 meshes of material granularity by using a ball mill, adding water and sulfuric acid to carry out mechanical stirring and continuous leaching, wherein the material-liquid ratio is 1 2 SO 4 The addition amount is 4 percent of the ore mass, the pH value of the ore pulp is controlled to be 1.4-2.5, the leaching time is 60min, and the total leaching rate of vanadium in the stone coal is 87.8 percent.
4. Solid-liquid separation and ion exchange: and (3) carrying out solid-liquid separation on the leached slurry through a belt type vacuum filter, carrying out fine pressing on the filtrate through a box type filter press, washing slag by using ion exchange tail water, and using washing water for the leaching process. After the refined pressure filtrate is clarified, the refined pressure filtrate is absorbed by anion resin D301, and tail water can be circulated to the processes of filtering, washing and leaching. Desorbing vanadium-containing saturated resin by using a desorbent (80 g/L NaOH solution) to obtain vanadium-containing saturated resin with the vanadium concentration of 82.6g V 2 O 5 L of enriched vanadium liquid.
5. Aging: adding sulfuric acid into the vanadium-enriched liquid to adjust the pH value to 9, standing and aging for 72h to precipitate a large amount of silicon dioxide precipitate, and filtering.
6. Deeply removing impurities: adding Ethylene Diamine Tetraacetic Acid (EDTA) accounting for 0.5% of the mass of the filtrate into the aged filtrate, stirring for 20 minutes, adding magnesium chloride accounting for 4% of the mass of the filtrate, uniformly stirring, and keeping the temperature at 70 ℃ for standing for 4 hours; cooling the vanadium solution to 40 ℃, performing filter pressing, and cooling the obtained filtrate in a vanadium precipitation reaction kettle;
7. and (3) vanadium precipitation: adding sulfuric acid into the filtrate in the vanadium precipitation reaction kettle to adjust the pH value to 1.8, heating the liquid to 90 ℃, stirring, and precipitating vanadium in the liquid in the form of ammonium polyvanadate. Filtering by a centrifuge to obtain solid high-purity ammonium polyvanadate. The vanadium precipitation tail water can be reused in the leaching process.
8. And (3) calcining: the high-purity ammonium polyvanadate is calcined in a rotary kiln at 550 ℃ for 4 hours, the calcined solid powder is brick red, namely a vanadium pentoxide product, the purity of the vanadium pentoxide product is up to 99.81% by analysis, and the high-purity vanadium pentoxide product meets the quality requirement of the electrolyte.
Example 2
1. Decarburization in a fluidized bed furnace: containing V 2 O 5 After 0.87wt% stone coal raw ore (calorific value 800-1200 Kcal/kg) is coarsely crushed by a jaw crusher,and then crushing the ore in a combined crusher, screening the ore by using a vibrating screen, controlling the ore particles to be less than 8mm, and returning unqualified materials through a belt conveyor. Adding lime with the sulfur content of 3 times of that of the ore into the crushed materials, detecting that the sulfur content of the ore is 0.68wt%, adding CaO with the sulfur content of 2wt%, uniformly mixing, feeding the mixture into a circulating fluidized bed boiler by a feeder to perform sulfur fixation, activation and roasting, wherein the roasting time is about 60min, the roasting temperature in the boiler is 900 ℃, and controlling the oxygen content in the smoke at the outlet of the boiler: and the volume percent is more than or equal to 5 percent, and the boiler flue gas is subjected to centralized treatment and then reaches the standard to be discharged.
2. Activating and roasting in a tunnel kiln: collecting the bottom slag and ash slag after sulfur fixation activation roasting in the circulating fluidized bed boiler, adding CaO accounting for 2 percent of the mass of the low slag and ash slag and BaSO accounting for 4 percent of the mass of the low slag and ash slag 4 Crushing the mixture to about 80 meshes by using a ball mill, making the brick by using a double-machine vacuum extruder, stacking the brick to a kiln car, and entering a tunnel kiln for high-temperature oxygen-enriched roasting. High temperature zone (temperature 900 ℃) residence time: 4 hours, the total retention time is 9 hours, the oxygen content in the flue gas is as follows: more than or equal to 10vol%, and the flue gas is discharged after reaching standards after being treated in a centralized way.
3. Continuous leaching of roasted clinker: crushing the roasted clinker to 60 meshes of material granularity by using a ball mill, adding water and sulfuric acid to carry out mechanical stirring and continuous leaching, wherein the material-liquid ratio is 1 2 SO 4 The addition amount is 4 percent of the ore mass, the pH value of the ore pulp is controlled to be 1.4-2.5, the leaching time is 60min, and the total leaching rate of vanadium in the stone coal is 87.8 percent.
4. Solid-liquid separation and ion exchange: and (3) carrying out solid-liquid separation on the leached slurry through a belt type vacuum filter, carrying out fine pressing on the filtrate through a box type filter press, washing slag by using ion exchange tail water, and using washing water for the leaching process. After the refined pressure filtrate is clarified, the refined pressure filtrate is absorbed by anion resin D301, and tail water can be circulated to the processes of filtering, washing and leaching. Desorbing the vanadium-containing saturated resin by using a desorbent (80 g/L NaOH solution) to obtain the vanadium-containing saturated resin with the vanadium concentration of 85.3g V 2 O 5 L of enriched vanadium liquid.
5. Aging: adding sulfuric acid into the vanadium-enriched liquid to adjust the pH value to 9, standing and aging for 72h to precipitate a large amount of silicon dioxide precipitate, and filtering.
6. Deeply removing impurities: adding ethylenediamine tetraacetic acid (EDTA) accounting for 0.2% of the mass of the filtrate into the aged filtrate, stirring for 20 minutes, then adding aluminum sulfate accounting for 4% of the mass of the filtrate and aluminum chloride accounting for 2% of the mass of the filtrate, stirring uniformly, and then keeping the temperature at 70 ℃ and standing for 4 hours; cooling the vanadium solution to 40 ℃, performing filter pressing, and cooling the obtained filtrate in a vanadium precipitation reaction kettle;
7. precipitating vanadium: adding sulfuric acid into the filtrate in the vanadium precipitation reaction kettle to adjust the pH value to 1.8, heating the liquid to 90 ℃, stirring, and precipitating vanadium in the liquid in the form of ammonium polyvanadate. And filtering by a centrifugal machine to obtain solid high-purity ammonium polyvanadate. The vanadium precipitation tail water can be reused in the leaching process.
8. And (3) calcining: the high-purity ammonium polyvanadate is calcined in a rotary kiln at 550 ℃ for 4 hours, the calcined solid powder is brick red, namely a vanadium pentoxide product, the purity of the vanadium pentoxide product is up to 99.61% by analysis, and the high-purity vanadium pentoxide product meets the quality requirement of the electrolyte.
Example 3
1. Decarburization in a fluidized bed furnace: containing V 2 O 5 0.87wt% of stone coal raw ore (with calorific value of 800-1200 kcal/kg) is coarsely crushed by a jaw crusher, then enters a combined crusher for crushing, is screened by a vibrating screen, and is controlled to have ore particles below 8mm, and unqualified materials are returned by a belt conveyor. Adding lime with the sulfur content of 3 times of that of the ore into the crushed materials, detecting that the sulfur content of the ore is 0.68wt%, adding CaO with the sulfur content of 2wt%, uniformly mixing, feeding the mixture into a circulating fluidized bed boiler by a feeder to perform sulfur fixation, activation and roasting, wherein the roasting time is about 60min, the roasting temperature in the boiler is 900 ℃, and controlling the oxygen content in the smoke at the outlet of the boiler: and the volume percent is more than or equal to 5 percent, and the boiler flue gas is subjected to centralized treatment and then reaches the standard to be discharged.
2. Activating and roasting in a tunnel kiln: collecting the bottom slag and ash slag after sulfur fixation activation roasting in the circulating fluidized bed boiler, adding CaO accounting for 2 percent of the mass of the low slag and ash slag and BaSO accounting for 4 percent of the mass of the low slag and ash slag 4 Crushing the mixture to about 80 meshes by using a ball mill, making the brick by using a double-machine vacuum extruder, stacking the brick to a kiln car, and entering a tunnel kiln for high-temperature oxygen-enriched roasting. Residence time in the high temperature zone (temperature 900 ℃): 4 hours, the total retention time is 9 hours, and the oxygen content in the flue gas is as follows: more than or equal to 10vol%, and the flue gas is subjected to centralized treatment and then reaches the standard to be discharged.
3. Continuous leaching of roasted clinker: pulverizing the calcined clinker by using a ball millAdding water and sulfuric acid to the material with the granularity of 60 meshes, mechanically stirring and continuously leaching, wherein the material-liquid ratio is 1 2 SO 4 The addition amount is 4 percent of the ore mass, the pH value of the ore pulp is controlled to be 1.4-2.5, the leaching time is 60min, the total leaching rate of vanadium in the stone coal is 87.8 percent, and the method is far higher than the conventional roasting vanadium extraction method.
4. Solid-liquid separation and ion exchange: and (3) carrying out solid-liquid separation on the leached slurry through a belt type vacuum filter, carrying out fine pressing on the filtrate through a box type filter press, washing slag by using ion exchange tail water, and using washing water for the leaching process. After the refined filtrate is clarified, the refined filtrate is absorbed by using anion resin D301, and tail water can be circulated to the processes of filtering, washing and leaching. Desorbing the vanadium-containing saturated resin by using a desorbent (80 g/L NaOH solution) to obtain the vanadium-containing saturated resin with the vanadium concentration of 81.2g V 2 O 5 L of enriched vanadium liquid.
5. Aging: adding sulfuric acid into the vanadium-enriched liquid to adjust the pH value to 9, standing and aging for 72h to precipitate a large amount of silicon dioxide precipitate, and filtering.
6. Deeply removing impurities: adding ethylenediamine tetraacetic acid (EDTA) accounting for 0.6% of the mass of the filtrate into the aged filtrate, stirring for 20 minutes, then adding aluminum sulfate accounting for 6% of the mass of the filtrate, uniformly stirring, and keeping the temperature at 70 ℃ and standing for 4 hours; cooling the vanadium solution to 40 ℃, performing filter pressing, and cooling the obtained filtrate in a vanadium precipitation reaction kettle;
7. precipitating vanadium: adding sulfuric acid into the filtrate in the vanadium precipitation reaction kettle to adjust the pH value to 1.8, heating the liquid to 90 ℃, stirring, and precipitating vanadium in the liquid in the form of ammonium polyvanadate. Filtering by a centrifuge to obtain solid high-purity ammonium polyvanadate. The vanadium precipitation tail water can be reused in the leaching process.
8. And (3) calcining: the high-purity ammonium polyvanadate is calcined in a rotary kiln at 550 ℃ for 4 hours, the calcined solid powder is brick red, namely a vanadium pentoxide product, the purity of the vanadium pentoxide product is up to 99.72% by analysis, and the high-purity vanadium pentoxide quality requirement of the electrolyte is met.
Claims (10)
1. The method for extracting vanadium from vanadium-containing stone coal is characterized by comprising the following steps:
(1) Roasting and leaching: roasting and leaching the vanadium-containing stone coal powder to obtain vanadium-containing leaching solution;
(2) Enrichment: enriching vanadium in the vanadium-containing leaching solution to obtain a vanadium-enriched solution;
(3) Aging: adjusting the pH value of the vanadium-enriched liquid to 8.5-9.5, standing, aging and filtering to obtain an aged filtrate;
(4) Removing impurities: and adding an impurity removing agent into the aged filtrate, and carrying out heat preservation, standing and filtering to obtain the vanadium-containing purified liquid, wherein the impurity removing agent comprises a chelating agent and a flocculating agent.
2. The method of claim 1,
in the step (1), the roasting comprises the steps of carrying out sulfur fixation activation roasting, and then adding a calcium compound and a barium compound into the roasted slag after the sulfur fixation activation roasting for roasting; and/or
In the step (1), the leaching includes soaking the roasted material with an acidic solution, and the acidic solution is preferably a sulfuric acid solution.
3. The method of claim 1,
in the step (2), the enrichment is carried out by adopting an adsorption mode of anion exchange resin, and the anion exchange resin is preferably D301 resin; and/or
In the step (2), the vanadium concentration of the vanadium-enriched liquid is more than or equal to 60gV 2 O 5 /L。
4. The method according to claim 3, wherein in the step (2), the anion exchange resin adsorbed with vanadium is desorbed using a lye.
5. The method of claim 1,
in the step (3), the pH value of the vanadium-enriched liquid is adjusted to 8.5-9.5 by adopting sulfuric acid; and/or
In the step (3), the standing and aging time is more than or equal to 48h.
6. The method of claim 1, wherein step (4) has one or more of the following characteristics:
adding impurity removing agent, keeping the temperature at 60-80 deg.C, and standing for 4-6 hr;
the chelating agent is ethylenediamine tetraacetic acid;
the flocculating agent is one or more selected from aluminum sulfate, magnesium chloride and aluminum chloride;
the dosage of the chelating agent is 0.1 to 1 percent of the mass of the filtrate;
the dosage of the flocculating agent is 1-10% of the mass of the filtrate;
adding the chelating agent, stirring uniformly, for example, stirring for 10-30min, and then adding the flocculating agent.
7. The method of claim 1, further comprising the steps of:
(5) And (3) vanadium precipitation: precipitating vanadium in the vanadium-containing purification solution in the form of ammonium polyvanadate, and carrying out solid-liquid separation to obtain the ammonium polyvanadate.
8. The method of claim 7, wherein in the step (5), the pH value of the vanadium-containing purification solution is adjusted to 1.5-2.1, the vanadium-containing purification solution is heated to 80-100 ℃, and the vanadium is stirred to precipitate in the form of ammonium polyvanadate.
9. The method of claim 7, further comprising the steps of:
(6) And (3) calcining: calcining the ammonium polyvanadate to decompose the ammonium polyvanadate to obtain vanadium pentoxide.
10. The method according to claim 9, wherein in the step (6), the calcination temperature is 500 to 550 ℃ and the calcination time is 3 to 5 hours.
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