CN114686704B - Combined smelting process of molybdenum ore and tungsten ore - Google Patents
Combined smelting process of molybdenum ore and tungsten ore Download PDFInfo
- Publication number
- CN114686704B CN114686704B CN202011630584.2A CN202011630584A CN114686704B CN 114686704 B CN114686704 B CN 114686704B CN 202011630584 A CN202011630584 A CN 202011630584A CN 114686704 B CN114686704 B CN 114686704B
- Authority
- CN
- China
- Prior art keywords
- ore
- molybdenum
- tungsten
- leaching
- raffinate
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Active
Links
Images
Classifications
-
- 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
- C22B34/30—Obtaining chromium, molybdenum or tungsten
- C22B34/36—Obtaining tungsten
-
- 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
-
- 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
-
- 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
- C22B34/30—Obtaining chromium, molybdenum or tungsten
- C22B34/34—Obtaining molybdenum
-
- 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 relates to a combined smelting process of molybdenum ore and tungsten ore, belongs to the technical field of nonferrous metal smelting, and solves the problems of difficult treatment of smelting wastewater and high smelting cost in the independent smelting process of tungsten-molybdenum ore in the prior art. The invention relates to a combined smelting process of molybdenum ore and tungsten ore, which is used for smelting molybdenum concentrate and scheelite in a combined way; the molybdenum concentrate is subjected to oxygen pressure leaching by a molybdenum ore leaching aid, and is filtered to obtain molybdenum ore filtrate, and the molybdenum ore filtrate is extracted to obtain molybdenum ore raffinate; and adopting the molybdenum ore raffinate as a tungsten ore leaching agent to perform normal pressure leaching of scheelite. Realizes clean smelting of tungsten-molybdenum ore and comprehensive utilization of resources.
Description
Technical Field
The invention relates to the technical field of nonferrous metal smelting, in particular to a combined smelting process of molybdenum ore and tungsten ore.
Background
More than 99% of molybdenum in nature exists in the form of molybdenite, which is the most important mineral raw material for extracting molybdenum. The existing molybdenite smelting process mostly adopts an oxidizing roasting-ammonia leaching process, and a large amount of low-concentration sulfur dioxide smoke is generated in the oxidizing roasting process, so that sulfur elements in the molybdenite cannot be effectively recycled, and a large amount of air pollution is caused. Meanwhile, the temperature control difficulty is high in the roasting process, and the temperature is too high in the roasting process, so that molybdenum can be sublimated and lost; if the temperature in the roasting process is too low, incomplete molybdenum oxidation can be caused, and leaching difficulty is increased.
The scheelite as a main smelting raw material of tungsten is mined year by year, but the scheelite has the conditions of reduced grade, complex components and accompaniment with other metals, so that the ore dressing and smelting cost of the tungsten resource is increased. The conventional method for treating tungsten ore is a high-pressure alkaline leaching method, and the process can decompose tungsten ore under the conditions of high alkali, high temperature and high pressure, but the tungsten content in slag is high, so that low-grade complex tungsten ore is difficult to treat. Meanwhile, for scheelite with higher barium content, the tungsten and the barium in the scheelite exist in a very stable barium tungstate form, so that the decomposition is very difficult, the higher the barium content is, the larger the influence on the decomposition is, the effective decomposition is difficult to be carried out by the traditional high-pressure alkaline leaching method, and the leaching rate of tungsten is low.
Chinese patent document CN104372169A discloses a method for extracting tungsten from high barium tungsten ore, which comprises mixing high barium tungsten ore with a certain amount of SiO 2 With Na and Na 2 CO 3 Uniformly mixing, roasting at high temperature, and using Na for roasting slag 2 CO 3 The solution leaches tungsten, and the leaching rate of tungsten is improved. The method can effectively extract tungsten from the high-barium tungsten ore, but in the roasting step, the temperature needs to reach 800-1000 ℃, the energy consumption is very high, the cost is high, and the industrial application of the method is limited.
At present, the acid method cooperative treatment process developed for scheelite not only can efficiently treat low-grade scheelite, but also can realize the leaching of the scheelite under normal pressure, has no sodium salt wastewater discharge, is environment-friendly, has very wide application prospect, and consumes only sulfuric acid. If the molybdenum pressure-cooking waste acid and the scheelite acid smelting technology can be organically combined in a proper way, the problems of molybdenum smelting waste acid output and tungsten smelting acid consumption can be effectively solved. However, due to the problems of the difference of smelting technical routes and the like, the combined smelting of tungsten ore and molybdenum ore cannot be realized.
Disclosure of Invention
In view of the above analysis, the present invention aims to provide a combined smelting process for molybdenum ore and tungsten ore, which is used for solving the problem that the existing smelting process for molybdenum ore and tungsten ore respectively independent from each other cannot comprehensively and effectively utilize the reagents used in smelting.
The invention provides a combined smelting process of molybdenum ore and tungsten ore, which is used for smelting molybdenum concentrate and scheelite in a combined way;
the molybdenum concentrate is subjected to oxygen pressure leaching by a molybdenum ore leaching aid, and is filtered to obtain molybdenum ore filtrate, and the molybdenum ore filtrate is extracted to obtain molybdenum ore raffinate;
and adopting the molybdenum ore raffinate as a tungsten ore leaching agent to perform normal pressure leaching of scheelite.
Further, the scheelite smelting is leached by a tungsten ore leaching agent, the leached tungsten ore filtrate is obtained by filtering, and the tungsten ore filtrate is extracted to obtain tungsten ore raffinate;
the molybdenum ore leaching aid is obtained by regenerating tungsten ore raffinate through impurity removal and regeneration agents. Further, the molybdenum ore leaching aid is phosphoric acid.
Further, the main components of the molybdenum concentrate are as follows in percentage by mass: molybdenum 40-50%, bismuth 2-5%.
Further, the main component of the scheelite is tungsten trioxide 40%.
Further, the oxygen pressure of the oxygen pressure digestion leaching is 0.8Mpa to 1.5Mpa.
Further, in the oxygen autoclaving leaching of the molybdenum concentrate, the usage amount of the leaching aid is 0.5 to 1.5 times of the mass of the molybdenum concentrate.
Further, the oxygen autoclaving time is 2 to 5 hours.
Further, the temperature of the oxygen autoclaving is 180 to 210 ℃.
Further, the molybdenum concentrate is leached by oxygen autoclaving, and the leaching rate of molybdenum is more than 99 percent.
Further, the molybdenum ore raffinate includes phosphoric acid and sulfuric acid.
Further, the liquid-solid ratio of the leaching of the tungsten ore leaching agent is 8mL/g to 12mL/g.
Further, the tungsten ore leaching agent is leached for 5 hours to 7 hours.
Further, the scheelite leaching agent is leached, and the leaching rate of tungsten is more than 98%.
Further, molybdenum ore filtrate is extracted to obtain molybdenum ore extract, and the molybdenum ore extract is subjected to back extraction, evaporation and crystallization by a molybdenum ore back extractant to obtain a molybdenum product.
Further, the extractant extracted from the molybdenum ore filtrate is a neutral phosphine cation extractant, and the neutral phosphine cation extractant is one or a combination of P204 and P507.
Further, the concentration of the extractant is 10-50% by mass percent.
Further, the extraction of molybdenum is compared to O/a=4:1 to 2:1.
Further, the molybdenum extraction mode is countercurrent extraction.
Further, the extraction stage number of the molybdenum is 4 to 7.
Further, the molybdenum ore stripping agent is hydrogen peroxide.
Further, the mass fraction of the hydrogen peroxide is 10-20%.
Further, the extraction of strip molybdenum is compared to O/a=7:1 to 5:1.
Further, the extraction mode of the back-extracted molybdenum is countercurrent extraction.
Further, the extraction level of the back-extracted molybdenum is 2 to 4.
Further, tungsten ore extraction liquid is obtained by tungsten ore filtrate extraction, and tungsten product is obtained by tungsten ore extraction liquid through back extraction, evaporation and crystallization of tungsten ore back extractant.
Further, the extractant for extracting the tungsten ore filtrate is TBP.
Further, the tungsten ore stripping agent is ammonia water solution.
Further, the concentration of the aqueous ammonia solution is 3mol/L to 8mol/L.
Further, the impurity removing agent is one or more of calcium oxide, calcium hydroxide, calcium phosphate and calcium carbonate.
Further, the regenerant is sulfuric acid.
Compared with the prior art, the invention has at least one of the following beneficial effects:
(1) According to the invention, phosphoric acid is used as an auxiliary leaching agent for oxygenic pressure boiling leaching to convert sulfur element in molybdenum concentrate into sulfuric acid, so that digestion and leaching of molybdenum concentrate ore are completed, and after molybdenum is extracted by a molybdenum ore extracting agent, mixed acid of sulfuric acid and phosphoric acid is continuously fully utilized for leaching tungsten ore, compared with the prior art, a large amount of leaching agent is not required to leach molybdenum concentrate, so that impurity elements in minerals are effectively utilized, and the leaching agent is directly used for leaching scheelite after target element molybdenum is extracted, so that reagents used in a process are greatly saved, and waste is changed into valuable;
(2) According to the invention, after the leaching of scheelite is completed by using sulfuric acid and phosphoric acid mixed acid, phosphoric acid is recovered through simple impurity removal and regeneration, and is reused in the oxygen pressure boiling leaching of molybdenum concentrate, so that the recycling of the leaching aid in the oxygen pressure boiling leaching of molybdenum concentrate is realized, the reagents used in the process are greatly saved, and clean smelting is realized;
(3) According to the invention, the sulfur in the impurity element in the molybdenum concentrate is oxidized into sulfuric acid, so that the leaching digestion of the molybdenum concentrate is realized, the leaching technology is used for scheelite, the phosphoric acid after scheelite leaching is used for oxygen pressure boiling leaching of the molybdenum concentrate after impurity removal and recovery, the elements in the molybdenum concentrate are fully utilized for scheelite smelting, and the phosphoric acid after tungsten ore smelting is fully utilized for molybdenum concentrate smelting, so that the combined smelting of tungsten ore and molybdenum ore is realized.
In the invention, the technical schemes can be mutually combined to realize more preferable combination schemes. Additional features and advantages of the invention will be set forth in the description which follows, and in part will be obvious from the description, or may be learned by practice of the invention. The objectives and other advantages of the invention may be realized and attained by the structure particularly pointed out in the written description and drawings.
Drawings
The drawings are only for purposes of illustrating particular embodiments and are not to be construed as limiting the invention, like reference numerals being used to refer to like parts throughout the several views.
FIG. 1 is a flow chart of a combined smelting process of tungsten ore and molybdenum ore.
Detailed Description
The invention provides a combined smelting process of molybdenum ore and tungsten ore, wherein the process flow chart is shown in figure 1, and molybdenum concentrate and scheelite are smelted in a combined way;
the molybdenum concentrate is subjected to oxygen pressure leaching by a molybdenum ore leaching aid, and is filtered to obtain molybdenum ore filtrate, and the molybdenum ore filtrate is extracted to obtain molybdenum ore raffinate; and adopting the molybdenum ore raffinate as a tungsten ore leaching agent to perform normal pressure leaching of scheelite.
Specifically, the main components of the molybdenum concentrate are as follows in percentage by mass: molybdenum 40% to 50%, bismuth 2% to 5%; the main component of scheelite is tungsten trioxide 40%.
Specifically, the molybdenum ore leaching aid is phosphoric acid.
Specifically, the molybdenum ore raffinate includes phosphoric acid and sulfuric acid.
The invention adopts a treatment method combining oxygen autoclaving and leaching, phosphoric acid and/or calcium phosphate serve as an auxiliary leaching agent to play roles in assisting leaching and ore digestion in a system, under the condition of oxygen, oxygen oxidizes molybdenum disulfide in molybdenum concentrate, sulfur in the molybdenum disulfide is oxidized into IV-valent sulfur and further oxidized into VI-valent sulfuric acid, and under the action of sulfuric acid generated by oxidizing sulfur in the molybdenum concentrate, molybdenum in the molybdenum concentrate is converted into soluble molybdenum acyl cations so as to be completely dissolved in a liquid phase to form a solution.
Through experimental study, molybdenum concentrate cannot be digested by using the leaching aid phosphoric acid and/or calcium phosphate alone; oxygen is independently introduced into the water environment for oxypoaching, and the ores cannot be oxidized and digested. Therefore, under the condition of oxygen, phosphoric acid and/or calcium phosphate are used as leaching aids to carry out oxygen autoclaving to realize digestion and leaching of molybdenum concentrate ore, so as to realize leaching of molybdenum. While bismuth is not leached, is totally enriched in residual leaching residues and can be used for recovery or directly sold.
As the oxygen autoclaving continues, molybdenum sulphide in the ore is continually oxidized and combined with phosphate by the combined action of oxygen and the leaching aid, converted to phosphomolybdic acid and sulfuric acid, and the acidity increases progressively. Under a strong acid environment, the reaction balance of the mutual conversion of the phosphomolybdic heteropolyacid and the molybdenum acyl cations continuously moves to the direction of generating the molybdenum acyl cations, and molybdenum in the ore is dissolved and converted into a state of the molybdenum acyl cations with good solubility. The phosphoric acid or calcium phosphate assists molybdenum to be converted into phosphomolybdic heteropolyacid in the process of ore digestion and leaching, and is released in the process of converting phosphomolybdic heteropolyacid into molybdenum acyl cations, thereby playing a role in assisting leaching in the process of ore digestion and leaching.
It should be noted that, sulfur in molybdenum concentrate is oxidized into IV sulfur in the process of leaching oxygen autoclaving, IV sulfur is further oxidized into VI sulfuric acid, the sulfuric acid and the leaching aid phosphoric acid work together to complete digestion and leaching of ore, meanwhile, phosphoric acid combines with molybdenum to form phosphomolybdic heteropolyacid, and the phosphomolybdic acid is further converted into molybdenum acyl cations under the strong acid condition provided by sulfuric acid. The molybdenum acyl cations are cation exchanged with protons in the cation extractant into an organic phase, and sulfuric acid and phosphoric acid remain in the raffinate. The extraction of acid-containing molybdenum ore raffinate is a very important problem, and the neutralization with alkali and then the discharge are very wasteful resources. According to the research, the method fully utilizes the sulfuric acid obtained by strengthening in the molybdenum concentrate, and the raffinate containing the sulfuric acid and the phosphoric acid can be directly used for leaching scheelite without treatment. The sulfur in the molybdenum concentrate is leached out of the molybdenum ore firstly through oxidative autoclaving, and then the scheelite is leached out, so that the full utilization of sulfur elements in the molybdenum concentrate is realized, the impurity sulfur elements in the molybdenum concentrate are fully used for leaching treatment of molybdenum and tungsten ores, waste is changed into valuable things, the smelting production efficiency is improved, the cost is greatly saved, and the combined smelting of the molybdenum ore and the tungsten ore is realized.
Specifically, the oxygen pressure is 0.8Mpa to 1.5Mpa.
Oxygen is a key factor playing an oxidation role in the oxygen autoclaving process, and because the oxygen autoclaving leaching process is a heterogeneous reaction of gas-solid-liquid, the general gas participation reaction mainly depends on the heterogeneous reaction of the interface of gas-solid and gas-liquid, and the reaction rate of the interface reaction of gas participation is slow, so that the oxygen autoclaving leaching efficiency is seriously affected. The oxygen is provided with enough pressure, so that the oxygen is better dissolved in the liquid, and a gas-solid-liquid heterogeneous reaction is generated at a heterogeneous reaction interface of the liquid and the solid, thereby greatly improving the efficiency of oxygen autoclaving leaching.
Specifically, in the molybdenum concentrate and leaching aid system, the usage amount of the leaching aid is 0.5 to 1.5 times of the mass of the molybdenum concentrate.
Specifically, the press-cooking time is 2 hours to 5 hours.
Specifically, the leaching oxygen autoclaving temperature is 180 ℃ to 210 ℃.
The longer the oxygen autoclaving is, the more thorough the reaction is, but at the same time too long autoclaving results in higher energy consumption, so that autoclaving times of 2 to 5 hours are chosen.
The higher the leaching oxygen autoclaving temperature is, the faster the autoclaving speed is, but at the same time, the higher energy consumption is caused, so that the production efficiency and the energy consumption of autoclaving are comprehensively considered, and the leaching oxygen autoclaving temperature is determined to be 180-210 ℃ according to the efficiency-cost ratio.
Specifically, the leaching rate of molybdenum is more than 99%.
Specifically, the liquid-solid ratio of the leaching of the tungsten ore leaching agent is 8mL/g to 12mL/g.
Specifically, the time for leaching the tungsten ore leaching agent is 5 to 7 hours.
Specifically, the scheelite leaching agent is used for leaching, and the leaching rate of tungsten is more than 98%.
Because the main component of the scheelite is tungsten trioxide, a large amount of acid is required to be consumed in the leaching process of the scheelite in the prior art, the acid consumed in the scheelite leaching of the application is derived from the oxygen autoclaving process of molybdenum concentrate, sulfur in the molybdenum concentrate is oxidized into sulfuric acid through oxygen oxidation, and the sulfuric acid is used for leaching the scheelite after the leaching of the molybdenum ore is realized, so that the combined smelting of the molybdenum ore and the tungsten ore is realized.
Specifically, a cationic extractant is used for extracting molybdenum from molybdenum leaching solution, so as to obtain molybdenum raffinate and a molybdenum-loaded cationic extractant.
Specifically, the cation extractant is one or a combination of P204 and P507, and the molybdenum stripping agent is hydrogen peroxide.
Specifically, the concentration of the cation extractant is 10-50% by mass percent.
Specifically, the extraction of molybdenum is compared to O/a=4:1 to 2:1.
The ratio of the extraction is an important factor of the extraction, and when the ratio of the O/A ratio of the extraction is less than 2:1, the organic extraction is insufficient, and part of molybdenum cannot be transferred into the organic phase, so that from the viewpoint of the recovery rate of molybdenum, the larger the ratio of the O/A, the more thoroughly the molybdenum is transferred into the organic phase, and the less residues remain in the aqueous phase. However, too much organic phase would also result in increased costs of solvents, processes, etc., and it has been studied that when O/a is greater than 4:1, the extraction efficiency/cost ratio is severely reduced, thus selecting an extraction of molybdenum to be extracted versus O/a=4:1 to 2:1.
Specifically, the molybdenum extraction mode is countercurrent extraction.
Specifically, the extraction stage number of molybdenum extraction is 4 to 7.
Specifically, molybdenum ore filtrate is extracted to obtain molybdenum ore extract, and the molybdenum ore extract is subjected to back extraction, evaporation and crystallization by a molybdenum ore back extractant to obtain a molybdenum product.
Specifically, the mass fraction of the hydrogen peroxide is 10-20%.
Specifically, the extraction of strip molybdenum is compared to O/a=7:1 to 5:1.
Specifically, the extraction mode of the back-extracted molybdenum is countercurrent extraction.
Specifically, the extraction level of the back-extracted molybdenum is 2 to 4.
The leaching solution is subjected to chemical extraction by the cation extractant, hydrogen ions in the cation extractant and leached molybdenum acyl cations are subjected to cation exchange, and the molybdenum acyl cations are transferred to an organic phase.
In the chemical equilibrium process of the molybdenum acyl cations in the organic phase, a small amount of molybdenum acyl cations enter the water phase and are converted into molybdate ions, in the back extraction process, a small amount of molybdic acid radicals which enter the water phase and are converted into peroxomonosylate anions by the hydrogen peroxide serving as a back extractant promote the chemical equilibrium to move towards the direction of the molybdenum acyl cations converted into molybdate ions, and all the molybdenum acyl cations are converted into peroxomonosylate anions which are thoroughly separated from the organic phase, so that the back chemical extraction of molybdenum from the molybdenum-loaded cationic extractant is realized, and all the molybdenum ions are back extracted into the water phase of the back extractant.
Specifically, the main component of the molybdenum concentrate comprises 45 to 55 percent of molybdenum in percentage by mass.
Specifically, the recovery rate of molybdenum is 99% or more.
Specifically, tungsten ore filtrate is extracted to obtain tungsten ore extract, and tungsten ore extract is subjected to back extraction, purification and crystallization to obtain tungsten products.
Specifically, the extracting agent for extracting the tungsten ore filtrate is TBP, and the tungsten ore back-extracting agent is ammonia water.
Specifically, the tungsten ore back-extraction agent is ammonia water solution.
Specifically, the concentration of the aqueous ammonia solution is 3mol/L to 8mol/L.
In one possible implementation, the scheelite smelting is leached by a tungsten ore leaching agent, the leached solution is filtered to obtain tungsten ore filtrate, and the tungsten ore filtrate is extracted to obtain tungsten ore raffinate; the molybdenum ore leaching aid is obtained by regenerating tungsten ore raffinate through impurity removal and regeneration agents.
Specifically, the impurity removing agent is one or more of calcium oxide, calcium hydroxide, calcium phosphate and calcium carbonate.
Specifically, the regenerant is sulfuric acid.
After the extraction of the tungsten ore leaching liquid is completed, a large amount of sulfuric acid is consumed, impurities generated in the tungsten ore leaching process are removed by adding the impurity removing agent, and meanwhile, part of phosphoric acid is converted into calcium dihydrogen phosphate, so that the calcium dihydrogen phosphate is regenerated by adding the regenerating agent sulfuric acid. And (3) through impurity removal and regeneration, the phosphoric acid in the raffinate obtained after extraction of the tungsten ore leaching solution is completely recovered and is circularly used as a leaching agent in the oxygen pressure boiling leaching process of the molybdenum ore, so that the recycling of the phosphoric acid is realized.
Preferred embodiments of the present invention will now be described in detail with reference to the accompanying drawings, which form a part hereof, and together with the description serve to explain the principles of the invention, and are not intended to limit the scope of the invention.
Example 1
In one embodiment of the invention, a combined smelting process of molybdenum ore and tungsten ore is disclosed, as shown in figure 1.
The main component of the molybdenum concentrate raw material is 45.1 percent of molybdenum and 2.59 percent of bismuth by mass percent.
Calcium phosphate is used as an auxiliary leaching agent, the dosage of the auxiliary leaching agent is 0.5 times of the mass of the molybdenum concentrate, the oxygen partial pressure is 0.8Mpa, and the temperature is 210 ℃ and the molybdenum concentrate is leached for 5 hours by oxygen autoclaving in an autoclave.
After the oxygen pressure boiling leaching is finished, cooling to room temperature and filtering to obtain leaching slag containing 16.1% of bismuth and leaching liquid containing molybdenum.
Wherein, the leaching rate of molybdenum is up to 99 percent.
For the leaching solution containing molybdenum, firstly, 50% of cationic extractant P204 is used for extracting molybdenum to obtain molybdenum-loaded P204, the extraction ratio is O/A=4:1, and the extraction stages are five countercurrent stages.
Carrying out back extraction on the P204 loaded with molybdenum by using 15% hydrogen peroxide to obtain molybdenum back extraction liquid and molybdenum raffinate; the back extraction phase ratio is O/A=7:1, and the back extraction stage number is countercurrent three.
The total recovery rate of molybdenum in the extraction back-extraction process reaches 99 percent.
And evaporating and crystallizing the molybdenum strip liquor to obtain molybdic acid, and calcining to obtain a molybdenum trioxide product.
Adding molybdenum raffinate into scheelite, wherein the liquid-solid ratio (ml/g) is 10:1, heating to 95 ℃, and leaching the scheelite for 5 hours, wherein the leaching rate of tungsten reaches 98%;
and after the tungsten ore leaching is finished, cooling to room temperature, and filtering to obtain leaching liquid containing tungsten.
For a leaching solution containing tungsten, firstly, 50% of TBP is used for extracting tungsten to obtain TBP loaded with tungsten, the extraction ratio is O/A=4:1, and the extraction stage number is five counter-current.
And carrying out back extraction on the TBP loaded with tungsten by using 5mol/L ammonia water solution to obtain tungsten back extraction liquid and tungsten raffinate.
Evaporating and crystallizing the tungsten back extraction liquid to obtain ammonium paratungstate.
The raffinate of tungsten is neutralized to pH value equal to 4.5 by calcium oxide, filtered to obtain solution containing monocalcium phosphate, the pH value is regulated to 1 by dilute sulfuric acid to obtain purified solution, and the purified solution is used as an infusion aid for autoclaving molybdenite and returns to be used, so that the internal circulation of leaching reagent is realized.
Example two
In one embodiment of the invention, a combined smelting process of molybdenum ore and tungsten ore is disclosed, as shown in figure 1.
The main component of the molybdenum concentrate raw material is 45.9% of molybdenum and 2.89% of bismuth by mass percent.
Phosphoric acid is used as an auxiliary leaching agent, the dosage of the auxiliary leaching agent is 0.8 times of the mass of molybdenum concentrate, the oxygen partial pressure is 1.5Mpa, and the temperature is 180 ℃ and the molybdenum concentrate is leached for 2 hours by oxygen autoclaving in an autoclave.
After the oxygen pressure boiling leaching is finished, cooling to room temperature and filtering to obtain leaching slag containing 16.5% of bismuth and leaching liquid containing molybdenum.
Wherein, the leaching rate of molybdenum is up to 99 percent.
For the leaching solution containing molybdenum, the molybdenum is firstly extracted by using 40% of cationic extractant P507 to obtain molybdenum-loaded P507, the extraction ratio is O/A=5:1, and the extraction stages are five countercurrent stages.
Back-extracting the P507 loaded with molybdenum by using 10% hydrogen peroxide to obtain molybdenum back-extraction liquid and molybdenum raffinate; the back extraction phase ratio is O/A=5:1, and the back extraction stage number is countercurrent three.
The total recovery rate of molybdenum in the extraction back-extraction process reaches 99 percent.
And evaporating and crystallizing the molybdenum strip liquor to obtain molybdic acid, and calcining to obtain a molybdenum trioxide product.
Adding scheelite into the molybdenum raffinate, leaching the scheelite for 5 hours with the liquid-solid ratio (ml/g) of 10:1, and heating to 95 ℃ until the leaching rate of the scheelite reaches 98.3%;
and after the tungsten ore leaching is finished, cooling to room temperature, and filtering to obtain leaching liquid containing tungsten.
For a leaching solution containing tungsten, firstly, 50% of TBP is used for extracting tungsten to obtain TBP loaded with tungsten, the extraction ratio is O/A=4:1, and the extraction stage number is five counter-current.
And carrying out back extraction on the TBP loaded with tungsten by using 5mol/L ammonia water solution to obtain tungsten back extraction liquid and tungsten raffinate.
Evaporating and crystallizing the tungsten back extraction liquid to obtain ammonium paratungstate.
The raffinate of tungsten is neutralized to pH value equal to 4.5 by calcium oxide, filtered to obtain solution containing monocalcium phosphate, the pH value is regulated to 1 by dilute sulfuric acid to obtain purified solution, and the purified solution is used as an infusion aid for autoclaving molybdenite and returns to be used, so that the internal circulation of leaching reagent is realized.
The present invention is not limited to the above-mentioned embodiments, and any changes or substitutions that can be easily understood by those skilled in the art within the technical scope of the present invention are intended to be included in the scope of the present invention.
Claims (7)
1. The combined smelting process of molybdenum ore and tungsten ore is characterized in that the molybdenum ore is molybdenum concentrate, and the tungsten ore is scheelite;
the molybdenum concentrate is subjected to oxygen pressure leaching by a molybdenum ore leaching aid, and molybdenum ore filtrate is obtained through filtration; wherein, the oxygen pressure of oxygen pressure cooking leaching is 0.8Mpa to 1.5Mpa, the dosage of the auxiliary leaching agent is 0.5 times to 1.5 times of the mass of the molybdenum concentrate in the oxygen pressure cooking leaching of the molybdenum concentrate, the oxygen pressure cooking time is 2 hours to 5 hours, and the temperature of the oxygen pressure cooking is 180 ℃ to 210 ℃; extracting the molybdenum ore filtrate to obtain molybdenum ore extract and molybdenum ore raffinate;
performing normal pressure leaching of scheelite by taking the molybdenum ore raffinate as a tungsten ore leaching agent;
leaching the scheelite by using molybdenum ore raffinate, filtering to obtain tungsten ore filtrate, and extracting the tungsten ore filtrate to obtain tungsten ore raffinate;
the tungsten ore raffinate is circularly used as a molybdenum ore leaching aid after impurity removal and regeneration; the molybdenum ore leaching aid is phosphoric acid, and the molybdenum ore raffinate comprises phosphoric acid and sulfuric acid.
2. The combined smelting process of molybdenum ore and tungsten ore according to claim 1, wherein the tungsten ore raffinate is purified by a purifying agent, wherein the purifying agent is one or more of calcium oxide, calcium hydroxide, calcium phosphate and calcium carbonate.
3. The combined smelting process of molybdenum ore and tungsten ore according to claim 1, wherein the tungsten ore raffinate is regenerated using a regenerant, the regenerant being sulfuric acid.
4. The combined smelting process of molybdenum ore and tungsten ore according to claim 1, wherein the molybdenum ore extract is back extracted by a molybdenum ore back extractant and evaporated and crystallized to obtain a molybdenum product.
5. The combined smelting process of molybdenum ore and tungsten ore according to claim 4, wherein the extractant extracted from the molybdenum ore filtrate is neutral phosphine cation extractant, and the molybdenum ore back extractant is hydrogen peroxide.
6. The combined smelting process of molybdenum ore and tungsten ore according to claim 1, wherein the tungsten ore filtrate is extracted to obtain tungsten ore extract, and the tungsten ore extract is back extracted by a tungsten ore back extractant and purified and crystallized to obtain tungsten product.
7. The combined smelting process of molybdenum ore and tungsten ore according to claim 6, wherein the extractant for extracting the tungsten ore filtrate is TBP and the tungsten ore stripping agent is ammonia water.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN202011630584.2A CN114686704B (en) | 2020-12-31 | 2020-12-31 | Combined smelting process of molybdenum ore and tungsten ore |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN202011630584.2A CN114686704B (en) | 2020-12-31 | 2020-12-31 | Combined smelting process of molybdenum ore and tungsten ore |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| CN114686704A CN114686704A (en) | 2022-07-01 |
| CN114686704B true CN114686704B (en) | 2023-05-30 |
Family
ID=82133421
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CN202011630584.2A Active CN114686704B (en) | 2020-12-31 | 2020-12-31 | Combined smelting process of molybdenum ore and tungsten ore |
Country Status (1)
| Country | Link |
|---|---|
| CN (1) | CN114686704B (en) |
Families Citing this family (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN115747530B (en) * | 2022-11-23 | 2023-12-01 | 郑州大学 | Method for separating and extracting molybdenum and tungsten from molybdenum-tungsten bulk concentrate |
Citations (8)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101323915A (en) * | 2008-07-17 | 2008-12-17 | 北京矿冶研究总院 | Method for extracting molybdenum and nickel from molybdenum-nickel ore by full wet method |
| CN102080161A (en) * | 2010-12-24 | 2011-06-01 | 中南大学 | Method for comprehensively recovering tungsten and phosphorus from high phosphorus white tungsten ores |
| CN103866142A (en) * | 2014-03-27 | 2014-06-18 | 西北有色金属研究院 | Method of recycling molybdenum and rhenium from molybdenum concentrate by hydrometallurgy |
| CN104928502A (en) * | 2015-05-13 | 2015-09-23 | 中南大学 | Method for extracting tungsten from wolframite or scheelite and wolframite mixture ore through adding of high-phosphorus scheelite |
| CN108396142A (en) * | 2018-05-03 | 2018-08-14 | 中南大学 | A kind of method of acid decomposed by phosphoric acid molybdenum calcining |
| CN108396141A (en) * | 2018-05-03 | 2018-08-14 | 中南大学 | A kind of clean metallurgical method of molybdenum |
| CN110564962A (en) * | 2019-10-14 | 2019-12-13 | 中南大学 | Smelting method of black-white tungsten mixed ore |
| CN111455172A (en) * | 2020-04-15 | 2020-07-28 | 中核沽源铀业有限责任公司 | Method for efficiently utilizing complex molybdenum ore by self-circulation of wastewater |
Family Cites Families (3)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US7485267B2 (en) * | 2005-07-29 | 2009-02-03 | Chevron U.S.A. Inc. | Process for metals recovery from spent catalyst |
| RU2532767C1 (en) * | 2010-12-24 | 2014-11-10 | Централ Саус Юниверсити | Method of extracting tungsten from scheelite |
| WO2013142022A1 (en) * | 2012-03-23 | 2013-09-26 | Kennecott Utah Copper Llc | Process for the conversion of molybdenite to molybdenum oxide |
-
2020
- 2020-12-31 CN CN202011630584.2A patent/CN114686704B/en active Active
Patent Citations (8)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101323915A (en) * | 2008-07-17 | 2008-12-17 | 北京矿冶研究总院 | Method for extracting molybdenum and nickel from molybdenum-nickel ore by full wet method |
| CN102080161A (en) * | 2010-12-24 | 2011-06-01 | 中南大学 | Method for comprehensively recovering tungsten and phosphorus from high phosphorus white tungsten ores |
| CN103866142A (en) * | 2014-03-27 | 2014-06-18 | 西北有色金属研究院 | Method of recycling molybdenum and rhenium from molybdenum concentrate by hydrometallurgy |
| CN104928502A (en) * | 2015-05-13 | 2015-09-23 | 中南大学 | Method for extracting tungsten from wolframite or scheelite and wolframite mixture ore through adding of high-phosphorus scheelite |
| CN108396142A (en) * | 2018-05-03 | 2018-08-14 | 中南大学 | A kind of method of acid decomposed by phosphoric acid molybdenum calcining |
| CN108396141A (en) * | 2018-05-03 | 2018-08-14 | 中南大学 | A kind of clean metallurgical method of molybdenum |
| CN110564962A (en) * | 2019-10-14 | 2019-12-13 | 中南大学 | Smelting method of black-white tungsten mixed ore |
| CN111455172A (en) * | 2020-04-15 | 2020-07-28 | 中核沽源铀业有限责任公司 | Method for efficiently utilizing complex molybdenum ore by self-circulation of wastewater |
Non-Patent Citations (1)
| Title |
|---|
| 十二磷钼酸的制备与开发;施友富;中国钼业(第04期);第13-15页 * |
Also Published As
| Publication number | Publication date |
|---|---|
| CN114686704A (en) | 2022-07-01 |
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| CN101967563B (en) | Method for wet-process vanadium extraction by using vanadium- and titanium-containing converter slag | |
| CN101265524B (en) | Method for extracting V2O5 from stone coal | |
| CN101760651B (en) | Process for extracting vanadium by acid leaching of stone coal | |
| CN102828025B (en) | Method for extracting V2O5 from stone coal navajoite | |
| CN102181635A (en) | Method for preparing vanadium pentoxide from stone coal vanadium ore sulfuric acid leaching solution | |
| CN107226484B (en) | A kind of method that compound extracting prepares vanadic anhydride | |
| CN104831075A (en) | Method for separating and purifying vanadium and molybdenum of waste vanadium-molybdenum SCR (selective catalytic reduction) catalyst | |
| CN104120269A (en) | Method for comprehensively utilizing vanadium slag | |
| CN105950865B (en) | A method of the separation and Extraction vanadium chromium from high-chromic vanadium leachate | |
| CN108866356A (en) | A method of the extraction vanadium extraction from vanadium slag high acidic oils liquid | |
| SE446015B (en) | PROCEDURE FOR THE RECOVERY OF COPPER BY LINING WITH SULFURIC ACID AND EXTRACTION WITH AMINE SOLVENT | |
| MX2013004855A (en) | Method for bioleaching and solvent extraction with selective recovery of copper and zinc from polymetal concentrates of sulfides. | |
| CN114686704B (en) | Combined smelting process of molybdenum ore and tungsten ore | |
| CN117327930B (en) | Method for recovering vanadium from primary shale stone coal | |
| Li et al. | Recovery of vanadium from direct acid leaching solutions of weathered crust vanadium‐titanium magnetite via solvent extraction with N235 | |
| CN115354173A (en) | Stone coal concentrated acid two-stage curing vanadium extraction process | |
| CN112320847B (en) | Method for extracting vanadium from high-phosphate vanadium solution and recycling wastewater | |
| CN111020233B (en) | Method for preparing vanadium pentoxide by ammonium-free vanadium precipitation | |
| CN111057875B (en) | Method for separating vanadium and chromium from solution by using microemulsion | |
| CN114686682B (en) | Comprehensive smelting method of molybdenite | |
| CN114686705B (en) | Method for recovering metal elements from molybdenite | |
| CN114686684B (en) | Method for recovering metal elements from molybdenum concentrate | |
| CN114686683B (en) | Method for recycling multiple metal elements from molybdenite based on molybdenum calcine | |
| KR101727891B1 (en) | Environment-Friendly Wet Process for Molybdenum Recovery From Spent Desulfurization Catalyst for Oil Refinery | |
| CN114686706B (en) | Method for recovering molybdenum and lead from molybdenum-lead ore |
Legal Events
| Date | Code | Title | Description |
|---|---|---|---|
| PB01 | Publication | ||
| PB01 | Publication | ||
| SE01 | Entry into force of request for substantive examination | ||
| SE01 | Entry into force of request for substantive examination | ||
| GR01 | Patent grant | ||
| GR01 | Patent grant |