CN101049968A - Method for associated producing ammonium molybdate and manganese sulfate by roasting molybdenite and pyrolusite - Google Patents

Abstract

The invention discloses a method for co-roasting molybdenite and pyrolusite to co-produce _{ammonium molybdate} and manganese _sulfate . b. Roast the mixture of the two together in the same equipment at a temperature of 350°C to 950°C for 60 to 360 minutes; c. Use inorganic acid to acidify the roasted product, and the acidolysis temperature is 5°C to 95°C, the liquid-solid ratio is 1.0-50.0, the acidolysis time is 5-300min, the pH value of the solution after acidolysis should be less than 5.5, and molybdenum salt, manganese salt solution and other insoluble impurities are obtained; d. The leaching solution is separated by solvent extraction Afterwards, obtain the extracting organic phase containing Mo and the raffinate aqueous phase containing Mn respectively; For extracting the organic phase containing Mo, adopt ammonium molybdate back-extraction to obtain ammonium molybdate solution; Ammonium molybdate and manganese sulfate products are produced correspondingly by metallurgical processes. The invention is a method for co-roasting molybdenite and pyrolusite to co-produce ammonium molybdate and manganese sulfate with short process, simple equipment, low energy consumption, low environmental pollution, high metal comprehensive recovery rate and high product quality.

Description

Method for co-roasting molybdenite and pyrolusite to co-produce ammonium molybdate and manganese sulfate

technical field

The invention relates to a method for directly producing ammonium molybdate and manganese sulfate by using molybdenite and pyrolusite.

Background technique

Molybdenum is a very important rare metal element, widely used in metallurgy, chemical industry, machinery, aerospace and biology and other fields, and has become an increasingly important metal material. my country has large reserves of molybdenum resources, wide distribution, many types, and many compound ores. The proven reserves are 4.15Mt, ranking second in the world. It is one of the countries with the largest export volume of molybdenum products in the world. The production capacity of molybdenum concentrate is rapid. increase, and has reached 70kt/a at present.

During the 1960s and 1970s, almost 90% of the production of pure molybdenum compounds in the world used the roasting-ammonia leaching process. The traditional roasting-ammonia leaching process is to oxidize and roast molybdenum concentrate at 600-700°C. Commonly used equipment includes shaft furnaces, reverberatory furnaces, mechanical furnaces (rakes), multi-hearth furnaces, fluidized fluidized furnaces, etc. The roasted molybdenum calcine (The main component is MoO ₃ ), ammonium molybdate products (I. Sasaki, US6559085; CJ Mahesh, US6190625) were obtained through steps such as ammonia leaching, purification and impurity removal, acid precipitation, filtration, and crystallization. MoO ₃ was significantly sublimated at 795°C. Due to the high temperature of the roasting process, the final recovery rate of Mo is only about 90%, and the energy consumption is high.

Lime-molybdenum concentrate co-roasting process and NaCO ₃ -molybdenum concentrate co-roasting process have been developed successively at home and abroad. The lime-molybdenum concentrate co-roasting process developed in the late 1980s is aimed at the recovery of SO ₂ pollution and associated rhenium elements during the roasting process of molybdenite. The main function of lime is to fix sulfur (CaSO ₄ ) and rhenium (Ca(ReO ₄ ) ₂ ), although the recovery rates of molybdenum and rhenium can be increased to 95% and 74% respectively, the problem of SO ₂ pollution has not been well solved, and the consumption of raw materials, waste residue and waste liquid in the production process Significantly increased, so it is difficult to popularize and apply (Liu Yinghan. Research on extracting molybdenum and rhenium from rhenium-containing molybdenum concentrate [J]. Journal of Jiangxi University (Natural Science Edition), 1989, 13(1): 88-95.) . The NaCO ₃ -molybdenum concentrate co-roasting process is mainly used to treat the ammonia leaching slag of molybdenum roasted sand with high molybdenum content. The principle is to convert the molybdate of Fe, Cu, Pb and other metals in the slag into more insoluble carbon salt, while molybdenum is converted into a sodium salt which is easily soluble in water.

At present, there are hundreds of industrial molybdenum oxide manufacturers in China, most of which adopt the traditional roasting-ammonia immersion process, but the overall level is relatively backward. Except for a few large-scale enterprises that use 8-12-layer multi-hearth furnaces, other small and medium-sized Manufacturers accounting for more than 1/3 of the total output use reverberatory furnaces or rotary kilns to roast molybdenum concentrates, and most of them do not have the means to recover sulfur dioxide, so they emit a large amount of low-concentration sulfur dioxide flue gas into the atmosphere every year. This situation has continued for the past 10 years.

Due to its stable properties, pyrolusite is insoluble in acid and alkali. The extraction of manganese requires conversion into acid-soluble low-priced Mn(II) oxides through reduction (adding a reducing agent), and then sulfuric acid is produced through sulfuric acid acidolysis and impurity removal. manganese products. At present, more than 30 domestic manufacturers adopt the reduction roasting-sulfuric acid leaching process to produce manganese sulfate, with a production capacity of 30,000 t/a, and the reduction roasting temperature requirement is as high as 700-1000°C. At the same time, a large amount of pulverized coal is consumed as a reducing agent. The process has high energy consumption, long process, and produces a large amount of CO ₂ and CO waste gas. The core step of extracting valuable metals from sulfide minerals (such as pyrite, sphalerite, and chalcopyrite) is to decompose metal sulfides through oxidation. Considering the strong oxidation of pyrolusite and the comprehensive utilization of resources, the above In the 1970s, the research on the two-mine leaching process with short process, simple equipment and low environmental pollution began (Mei Guanggui, Zhong Zhuqian. Simultaneous leaching of zinc sulfide concentrate and pyrolusite and simultaneous electrolysis of Zn-Mn Research [J]. Journal of Zhongnan Institute of Mining Management. 1982, (1): 18-25.), part of which has been successfully pushed to industrial production.

Contents of the invention

The technical problem to be solved by the present invention is to provide a molybdenite and pyrolusite co-roasted to co-produce ammonium molybdate and sulfuric acid with short process, simple equipment, low energy consumption, low environmental pollution, high metal comprehensive recovery rate and high product quality. manganese method.

In order to solve the problems of the technologies described above, the method that molybdenite and pyrolusite are jointly roasted and co-produced ammonium molybdate and manganese sulfate provided by the invention comprises the following steps:

a, molybdenite and pyrolusite are mixed according to the ratio of n _MnO2 : n _MoS2 = 1.0～10.0;

b. Co-calcining the mixture of the two in the same equipment at a temperature of 350°C to 950°C for 60 to 360 minutes;

c. Use inorganic acid to acidolyze the roasted product. The acidolysis temperature is 5°C-95°C, the liquid-solid ratio is 1.0-50.0, the acidolysis time is 5-300min, and the pH value of the solution after acidolysis should be less than 5.5 to obtain molybdenum salt , manganese salt solution and other insoluble impurities;

d, after the leaching solution is separated by solvent extraction, obtain the Mo-containing extraction organic phase and the Mn-containing raffinate aqueous phase; for the Mo-containing extraction organic phase, use ammonia water to back-extract to obtain ammonium molybdate solution;

e. The obtained ammonium molybdate solution and manganese salt solution are used to obtain ammonium molybdate and manganese sulfate products correspondingly by following the existing metallurgical process.

In the above step (a), the ratio of molybdenite to pyrolusite n _MnO2 : n _MoS2 = 2.0-5.0, the mixing of the two minerals is simple mechanical stirring mixing, or adding a binder to press molding.

In the above step (b), the two mineral mixtures are roasted in the same equipment, and there is no special requirement for the roasting equipment. The roasting equipment is a reverberatory furnace, a rotary kiln, a multi-hearth furnace or a fluidized bed roasting equipment in the current industrial production. Its production method It can be continuous production or batch production.

The calcining temperature in the above step (b) is 450°C-600°C.

The calcining atmosphere in the above step (b) is air or pure oxygen.

The inorganic acid used in the above step (c) is sulfuric acid, hydrochloric acid or nitric acid, and the acidolysis process needs to maintain a certain acidity of the system.

The extractant used in the above step (d) is a secondary amine. The extractant is N235, and other components of the extracted organic phase include sulfonated kerosene and secondary octanol.

The metallurgical process for recovering Mo from the ammonium molybdate solution in the above step (e) is purification, acid precipitation, and drying. When the amount of impurities is high, it is necessary to perform multi-step ammonia dissolution, acid precipitation and other processes.

The metallurgical process for recovering Mn from the Mn-containing solution in the above step (e) is to adjust the pH value, purify and remove impurities, evaporative crystallization, and recrystallization.

The method of the present invention is applicable to molybdenite of any grade, including molybdenite concentrate, middle ore or raw ore, etc.; the pyrolusite used can be concentrated ore after enrichment, or raw ore or middle ore of higher grade wait.

In the method of the present invention, the inorganic acid used can be sulfuric acid, hydrochloric acid, nitric acid, etc., and the acidolysis products are MoO ₂ SO ₄ , MoO ₂ Cl ₂ , MoO ₂ (NO ₃ ) ₂ solutions and manganese salts, such as MnSO ₄ , MnCl ₂ , Mn(NO ₃ ) ₂ solution.

The process of the present invention is generally carried out under normal pressure.

The method of co-roasting molybdenite and pyrolusite to co-produce ammonium molybdate and manganese sulfate by adopting the above technical scheme, after mixing molybdenite and pyrolusite in a certain proportion, they are fully roasted in the roasting equipment, and the sand product adopts X -Crystal diffraction analysis, confirmed that manganese molybdate is the main source; molybdenum salt can be obtained by acid hydrolysis of manganese-containing pebble sand, the acid hydrolysis is mainly inorganic acid, which can be sulfuric acid, hydrochloric acid, nitric acid, etc., and the acid hydrolysis product is MoO ₂ SO _4. MoO ₂ Cl ₂ , MoO ₂ (NO ₃ ) ₂ solutions and manganese salts, such as MnSO ₄ , MnCl ₂ , Mn(NO ₃ ) ₂ solutions; according to the occurrence state of Mo and Mn ions, use solvent extraction technology to separate The separation, purification and enrichment of Mo and Mn can obtain solutions containing Mo and Mn respectively, and then follow the existing process to produce ammonium molybdate and manganese sulfate products. The present invention is based on the strong oxidation of pyrolusite, the reducibility of molybdenite and the stability of manganese sulfate, and utilizes the new process of co-roasting molybdenite-pyrolurite to partially overcome the serious sulfur-containing flue gas in the traditional roasting-ammonia leaching molybdenum process. Pollution of the environment, improvement of metal comprehensive recovery rate and product quality (Mo recovery rate reaches more than 95%), and significant reduction of raw material consumption and energy consumption in pyrolusite decomposition process (such as high temperature reduction roasting-acid leaching process). The roasting process of the present invention has no special requirements on equipment and can be carried out in common molybdenum smelters.

In summary, the present invention is a short flow process, simple equipment, low energy consumption, low environmental pollution, high metal comprehensive recovery, high product quality molybdenite and pyrolusite joint roasting co-production of ammonium molybdate and manganese sulfate Methods.

Detailed ways

The invention is further illustrated by, but not limited to, the following examples. All parts and percentages in the examples refer to mass unless otherwise specified.

Example 1:

Stir and mix 5 parts of molybdenite concentrate (containing Mo 47.4%) and 8 parts of pyrolusite (containing Mn 50.4%) (n _MnO2 /n _MoS2 = 3.0), put the ore powder into the quartz tube and place it in the tube Heat up and roast in a type resistance furnace, connect one end of the quartz tube to a blower, control the air volume to 25ml/min, use a temperature controller to control the temperature at 500°C, and roast for 180min. Sulfur element analysis and Mo phase analysis were carried out on the calcined product, and the results showed that the sulfur fixation rate was 94.8%, the Mo oxidation rate was 100%, and the main component of the sand was manganese molybdate (MnMoO ₄ ).

The obtained manganese-containing Mo bei sand was leached with 100g/L sulfuric acid solution at room temperature (25°C), the solid ratio of the leaching liquid was 10, and the acid hydrolysis time was 90min, and the obtained manganese-containing Mo(MoO ₂ SO ₄ ), Mn(MnSO ₄ ) red solution and acid insoluble slag, the leaching rate of Mo is 99.8%, and the leaching rate of Mn is 89.0%. Use N235 (the extraction organic phase is 20% N235+10% 2-octanol+70% sulfonated kerosene) as the extractant to extract and recover Mo from the acid leaching solution containing Mn molybdenum calcine, the extraction temperature is room temperature (25 ° C), and the phase The ratio O/W is 1/5, the number of cross-flow extraction stages is 5, and the concentration of sulfuric acid in the water phase is ≤100g/L; when back-extracting, the extracted organic phase is first washed with dilute ammonia water, and the back-extraction agent uses 17% ammonia water , the stripping temperature is also at room temperature, compared to O/W being 1/2 and the extraction stage being 2, the extraction rate and stripping rate of Mo under this condition reach 99.6% and 99.1% respectively.

After the extraction process, ammonium molybdate solution and manganese sulfate solution (water raffinate phase) can be obtained respectively, and then the ammonium molybdate and manganese sulfate products are produced respectively according to the existing metallurgical process conditions, that is, ammonium sulfide is removed from the ammonium molybdate solution -Operations such as acid precipitation-ammonia dissolution-acid precipitation-drying, obtain the ammonium molybdate product that meets the national quality standard (national standard GB3460-82), the yield of Mo in this process is 96.9%, and the final recovery rate of Mo is 95.5%; Manganese sulphate solution also obtains the manganese sulphate product that meets the national quality standard through adjusting the pH value-purification and impurity removal-evaporative crystallization-recrystallization, etc., and the yield of Mn in this process is 90.3%, and the final recovery rate of Mn is 80.4%.

Example 2:

Stir and mix 5 parts of molybdenite concentrate (containing Mo 47.4%) and 16 parts of pyrolusite (containing Mn 50.4%) (n _MnO2 /n _MoS2 = 6.0), put the ore powder into the quartz tube and place it in the tube Heat up and roast in a type resistance furnace, connect one end of the quartz tube to a blower, control the air volume to 25ml/min, use a temperature controller to control the temperature at 600°C, and roast for 90 minutes. Sulfur element analysis and Mo phase analysis were carried out on the calcined product, and the results showed that the sulfur fixation rate was 98.8%, the Mo oxidation rate was 100%, and the main components of the sand were manganese molybdate (MnMoO ₄ ) and manganese sulfate.

The obtained manganese-containing Mo sand was leached with 300g/L sulfuric acid solution, the leaching was controlled at 65°C, the solid ratio of the leaching solution was 5, and the acid hydrolysis time was 300min, and the Mo(MoO ₂ SO ₄ ), Mn(MnSO ₄ ) The red solution and acid insoluble slag, the leaching rate of Mo is 99.8%, and the leaching rate of Mn is 73.4%. Use N235 (the extraction organic phase is 20% N235+10% 2-octanol+70% sulfonated kerosene) as the extractant to extract and recover Mo from the acid leaching solution containing Mn molybdenum calcine, the extraction temperature is room temperature (25 ° C), and the phase The ratio O/W is 1/5, the number of cross-flow extraction stages is 5, and the concentration of sulfuric acid in the water phase is ≤200g/L; when back-extracting, the extracted organic phase is first washed with dilute ammonia water, and the back-extraction agent uses 17% ammonia water , the stripping temperature is also at room temperature, compared to O/W being 1/2 and the extraction stage being 2, the extraction rate and stripping rate of Mo under this condition reach 99.6% and 98.2% respectively.

After the extraction process, ammonium molybdate solution and manganese sulfate solution (water raffinate phase) can be obtained respectively, and then the ammonium molybdate and manganese sulfate products are produced respectively according to the existing metallurgical process conditions, that is, ammonium sulfide is removed from the ammonium molybdate solution -Operations such as acid precipitation-ammonia dissolution-acid precipitation-drying, obtain the ammonium molybdate product that meets the national quality standard (national standard GB3460-82), the yield of Mo in this process is 96.9%, and the final recovery rate of Mo is 94.6%; Manganese sulphate solution also obtains the manganese sulphate product that meets the national quality standard through adjusting pH value-purification and impurity removal-evaporative crystallization-recrystallization etc. equally, the yield of this process Mn is 90.9%, and the final recovery rate of Mn is 66.7%.

Example 3:

Stir and mix 15 parts of molybdenite concentrate (containing Mo 47.4%) and 8 parts of pyrolusite (containing Mn 50.4%) (n _MnO2 /n _MoS2 = 2.0), put the ore powder into the quartz tube and place it in the tube Heat up and roast in a type resistance furnace, connect one end of the quartz tube to a blower, control the air volume to 45ml/min, use a temperature controller to control the temperature at 550°C, and roast for 120min. Sulfur element analysis and Mo phase analysis were carried out on the calcined product, and the results showed that the sulfur fixation rate was 91.8%, the Mo oxidation rate was 100%, and the main component of the sand was manganese molybdate (MnMoO ₄ ).

The obtained manganese-containing Mo sand is leached by 30% nitric acid solution, the leaching is room temperature (25° C.), the leaching liquid-solid ratio is 15, and the leaching time is 30 min, and the red solution and acid-insoluble slag containing Mo and Mn are obtained. The leaching rate was 99.9%, and the leaching rate of Mn was 98.2%. Subsequent steps were the same as in Example 1, the final recovery rate of Mo was 96.0%, and the final recovery rate of Mn was 91.6% (Mn(NO ₃ ) ₂ ).

Example 4:

Stir and mix 113 parts of molybdenite medium ore (containing Mo 26.4%) and 100 parts of pyrolusite (containing Mn 50.4%) (n _MnO2 /n _MoS2 = 3.0), put the ore powder into the quartz tube and place it in the tube Heat up and roast in a type resistance furnace, one end of the quartz tube is connected to a blower, the air volume is controlled to be 20ml/min, the temperature is controlled at 450 by a temperature controller, and roasted for 360min. Sulfur element analysis and Mo phase analysis were carried out on the roasted product, and the results showed that the sulfur fixation rate was 92.9%, the Mo oxidation rate was 98.7%, and the main component of the sand was manganese molybdate (MnMoO ₄ ).

The obtained manganese-containing Mo sand is leached by 17% hydrochloric acid solution, the leaching is room temperature (25°C), the leaching liquid-solid ratio is 30, and the leaching time is 180min, and the red solution and acid-insoluble slag containing Mo and Mn are obtained. The leaching rate was 97.9%, and the leaching rate of Mn was 88.2%. Subsequent steps were the same as in Example 1, the final recovery rate of Mo was 93.7%, and the final recovery rate of Mn was 80.5% (MnCl ₂ ).

Example 5:

Stir and mix 5 parts of molybdenite concentrate (containing Mo 47.4%) and 16 parts of pyrolusite raw ore (containing Mn 24.8%) (n _MnO2 /n _MoS2 = 3.0), put the ore powder into a quartz tube and place Heat up and roast in a tubular resistance furnace, connect one end of the quartz tube to a blower, control the air volume to 30ml/min, use a temperature controller to control the temperature at 550°C, and roast for 180min. Sulfur element analysis and Mo phase analysis were carried out on the calcined product, and the results showed that the sulfur fixation rate was 95.9%, the Mo oxidation rate was 100%, and the main component of the sand was manganese molybdate (MnMoO ₄ ). The follow-up processing steps of the sandstone are the same as in Example 1. The final recovery of Mo was 95.7%, and the final recovery of Mn was 91.2% (MnSO ₄ ).

Embodiment 6:

Stir and mix 5 parts of molybdenite concentrate (containing Mo 47.4%) and 16 parts of pyrolusite raw ore (containing Mn 24.8%) (n _MnO2 /n _MoS2 = 3.0), put the ore powder into a quartz tube and place Heat up and roast in a tubular resistance furnace, connect one end of the quartz tube to an oxygen bottle, control the oxygen flow rate to 10ml/min, use a temperature controller to control the temperature at 450°C, and roast for 150min. Sulfur element analysis and Mo phase analysis were carried out on the calcined product, and the results showed that the sulfur fixation rate was 93.5%, the Mo oxidation rate was 100%, and the main component of the sand was manganese molybdate (MnMoO ₄ ). The follow-up processing steps of the sandstone are the same as in Example 1. The final recovery of Mo was 95.9%, and the final recovery of Mn was 90.1% (MnSO ₄ ).

Embodiment 7:

5 parts of molybdenite concentrate (containing Mo 47.4%) and 16 parts of pyrolusite raw ore (containing Mn 24.8%) were stirred and mixed evenly (n _MnO2 /n _MoS2 = 3.0), and then placed in a muffle furnace for heating and roasting. The controller controls the temperature at 350°C for 60 minutes, and then increases the temperature to 450°C for 60 minutes. The Mo phase analysis of the calcined product shows that the oxidation rate of Mo is 99.8%, and the main component of the bei sand is manganese molybdate (MnMoO ₄ ). The follow-up processing steps of the sandstone are the same as in Example 1. The final recovery of Mo was 95.0%, and the final recovery of Mn was 89.5% (MnSO ₄ ).

Embodiment 8:

5 parts of molybdenite concentrate medium ore (containing Mo 26.4%) and 15 parts of pyrolusite raw ore (containing Mn 24.8%) were stirred and mixed evenly (n _MnO2 /n _MoS2 = 5.0) and then placed in a muffle furnace for heating and roasting. Use a temperature controller to control the temperature at 550° C., and bake for 100 minutes. Sulfur element analysis and Mo phase analysis were carried out on the roasted product, and the results showed that the sulfur fixation rate was 96.9%, the Mo oxidation rate was 100%, and the main component of the sand was manganese molybdate (MnMoO ₄ ). The follow-up processing steps of the sandstone are the same as in Example 1. The final recovery of Mo was 95.3%, and the final recovery of Mn was 74.9% (MnSO ₄ ).

Embodiment 9:

After stirring and mixing 5 parts of molybdenite concentrate medium ore (containing Mo 26.4%) and 6 parts of pyrolusite raw ore (containing Mn 24.8%) (n _MnO2 /n _MoS2 = 2.0), use 0.1% CMC solution as a viscous As a binder, the ore powder is pressed into a cylindrical shape (diameter 5mm, height 8mm), placed in a muffle furnace to raise the temperature and roasted, using a temperature controller to control the temperature at 700°C, and roasted for 240 minutes. The Mo phase analysis of the calcined product shows that the oxidation rate of Mo is 99.4%, and the main component of the sand is manganese molybdate (MnMoO ₄ ). The follow-up processing steps of the sandstone are the same as in Example 1. The final recovery of Mo was 94.7%, and the final recovery of Mn was 85.2% (MnSO ₄ ).

Example 10:

Stir and mix molybdenite concentrate medium ore (containing Mo 26.4%) and pyrolusite raw ore (containing Mn 24.8%) at a mass ratio of 5:3 (n _MnO2 /n _MoS2 = 1.0), and then continuously add it to the small rotary In the kiln, and continuously discharge the material; the temperature in the middle of the rotary kiln is controlled at 700-750°C, and the roasting time of the material is (residence time) 240min. The Mo phase analysis of the calcined product shows that the oxidation rate of Mo is 99.4%, and the main component of the sand is manganese molybdate (MnMoO ₄ ). The follow-up processing steps of the sandstone are the same as in Example 1. The final recovery of Mo was 96.3%, and the final recovery of Mn was 87.1% (MnSO ₄ ).

Claims (10)

1, the method for a kind of molybdenum glance and pyrolusite common roasting coproduction ammonium molybdate and manganous sulfate, it comprises the following step:

A, molybdenum glance and pyrolusite are pressed n _MnO2: n _MoS2=1.0～10.0 ratio mixings;

B, be 350 ℃～950 ℃ condition under common roasting in temperature with both mixture in same equipment, the time is 60～360min;

C, utilize mineral acid acidolysis product of roasting, acidolysis temperature is 5 ℃～95 ℃, and liquid-solid ratio is 1.0～50.0, and the acidolysis time is 5～300min, and the pH value of solution should obtain molybdenum salt, manganese salt solution and other insoluble impurities slag less than 5.5 after the acidolysis;

D, leach liquor after solvent extraction and separation, the raffinate water that obtains respectively containing the Mo extracted organic phase He contain Mn; For containing the Mo extracted organic phase, adopt the ammoniacal liquor back extraction to get ammonium molybdate solution;

E, the ammonium molybdate solution that obtains, manganese salt solution are continued to use existing corresponding ammonium molybdate and the manganese sulfate product of making of metallurgical technology.

2, the method for molybdenum glance according to claim 1 and pyrolusite common roasting coproduction ammonium molybdate and manganous sulfate is characterized in that: the ratio n of molybdenum glance and pyrolusite in the above-mentioned steps (a) _MnO2: n _MoS2The mixing of=2.0～5.0, two kinds of mineral is that simple mechanical stirring is mixed, or adds binding agent compression moulding.

3, the method for molybdenum glance according to claim 1 and 2 and pyrolusite common roasting coproduction ammonium molybdate and manganous sulfate, it is characterized in that: two kinds of mineral intermixture roastings in same equipment in the above-mentioned steps (b), its roasting apparatus is reverberatory furnace, rotary kiln, multiple hearth furnace or sulfuration bed roasting apparatus, its mode of production can be a continuous production, also can batch production.

4, the method for molybdenum glance according to claim 1 and 2 and pyrolusite common roasting coproduction ammonium molybdate and manganous sulfate is characterized in that: maturing temperature is 450 ℃～600 ℃ in the above-mentioned steps (b).

5, the method for molybdenum glance according to claim 1 and 2 and pyrolusite common roasting coproduction ammonium molybdate and manganous sulfate is characterized in that: calcination atmosphere is an air in the above-mentioned steps (b), or pure oxygen.

6, the method for molybdenum glance according to claim 1 and pyrolusite common roasting coproduction ammonium molybdate and manganous sulfate is characterized in that: used mineral acid is sulfuric acid, hydrochloric acid or nitric acid in the above-mentioned steps (c), and the acidolysis process need keep the certain acidity of system.

7, the method for molybdenum glance according to claim 1 and pyrolusite common roasting coproduction ammonium molybdate and manganous sulfate is characterized in that: used extraction agent is a secondary amine in the above-mentioned steps (d).

8, the method for molybdenum glance according to claim 7 and pyrolusite common roasting coproduction ammonium molybdate and manganous sulfate, it is characterized in that: described extraction agent is N235, other composition of extracted organic phase has sulfonated kerosene, secondary octanol.

9, the method for molybdenum glance according to claim 1 and pyrolusite common roasting coproduction ammonium molybdate and manganous sulfate, it is characterized in that: the metallurgical technology from ammonium molybdate solution recovery Mo in the above-mentioned steps (e) is that purification, acid are heavy, dry, and impurity level Gao Shixu carries out operations such as multistep ammonia is molten, acid is sunk.

10, the method for molybdenum glance according to claim 1 and pyrolusite common roasting coproduction ammonium molybdate and manganous sulfate is characterized in that: be to adjust pH value, purification and impurity removal, evaporative crystallization, recrystallization from containing the metallurgical technology that Mn solution reclaims Mn in the above-mentioned steps (e).