JP2015183217A - separation method - Google Patents

Abstract

PROBLEM TO BE SOLVED: To provide a separation method capable of separating mineral concentrate into mineral concentrate having a high copper content concentration and mineral concentrate having a high molybdenum concentration.SOLUTION: A separation method includes a heating step and a separation step. In the heating step, rough concentrate obtained by pulverizing and selecting yellow pyrite crude ore containing molybdenum is mixed with an elemental sulfur and then heated to 350°C-450°C. In the separation step, the rough concentrate obtained in the heating step is crushed to perform floatation method so as to be separated into first flotation concentrate recovered at a preceding stage of a process of the floatation method, and second floatation concentrate recovered at a subsequent stage of the process.

Description

  The present invention relates to a separation method.

The source of molybdenite (MoS ₂ ), which is a raw material for smelting molybdenum, is rarely produced alone, and is often collected as an accompaniment of galena (PbS) or copper sulfide ore. The quality of molybdenum as an accompanying material is about 50 to 500 ppm by weight as crude ore.

An ore containing copper sulfide ore and hydropyrite ore is finely crushed and then recovered as copper concentrate containing pyroxenite by a flotation process. Bright lead ore is known as one of the minerals that are likely to surface, and floats together with copper sulfide ore. The copper concentrate containing the molybdenite is further used for preferential flotation of molybdenite. As a result, the molybdenite can be recovered. In the preferential flotation of molybdenite, it is possible to concentrate the molybdenite ore by repeatedly performing flotation using a copper sulfide ore inhibitor and the like. Separation of molybdenite and copper sulfide ore, especially separation of chalcopyrite (CuFeS ₂ ), which is the most common mineral, is difficult. Is included.

  The bright lead ore concentrate is used as a lubricant raw material as molybdenum sulfide, or roasted into molybdenum oxide and used as a metallurgical raw material. However, as described above, it is difficult to separate the copper sulfide ore from the hydropyrite. In addition, the increase in the amount of copper in the hydrous ore concentrate will deteriorate the profits in the copper concentrate and significantly reduce the commercial value of the hydrous ore. In addition, a copper removal process for the bright lead ore concentrate is also required.

  The copper sulfide ore that is most accompanied by hydropyrite is chalcopyrite. Even if the chalcopyrite is subjected to flotation, the separation efficiency is low. Moreover, chalcopyrite is a chemically very stable mineral, and the method of selectively leaching and removing only chalcopyrite is extremely limited. Therefore, Patent Document 1 discloses a method of warm leaching with an oxidizing agent such as high concentration iron (III) chloride or ozone.

JP-A-5-195106

  However, the technique of Patent Document 1 increases the cost.

  In view of the above problems, an object of the present invention is to provide a separation method capable of separating a concentrate into a concentrate having a high copper concentration and a concentrate having a high molybdenum concentration while suppressing cost.

  The separation method according to the present invention is obtained by a heating process in which a coarse ore obtained by pulverizing and selecting a chalcopyrite ore containing molybdenum is mixed with elemental sulfur and heated to 350 ° C. to 450 ° C., and obtained in the heating process. The coarse concentrate is crushed and subjected to flotation, and the first flotation concentrate recovered in the previous stage of the flotation process and the second flotation concentrate recovered in the latter stage of the process are separated. And a separating step.

  The first flotation concentrate may be subjected to a copper smelting process, and the second flotation concentrate may be subjected to a molybdenum smelting process. The chalcopyrite coarse ore may contain molybdenite. The chalcopyrite coarse ore may contain at least one of pyrite, chalcocite, and copper indigo. The coarse beneficiation may have a copper grade of 3 to 15% by weight and an 80% passing particle size of 100 to 300 μm. In the heating step, 1.0 to 2.0 times by weight of simple sulfur may be added to the copper content in the coarse beneficiation. In the separation step, the coarse concentrate may be crushed until the 80% passing particle size becomes 30 to 100 μm, and then the flotation may be performed. In the separation step, after recovering the first floating concentrate, the collection material is added to the floating beneficiation tank in an amount of 100 g / t or more, and the pH of the tank liquid is adjusted to 10.0 to 12.5. You may collect | recover as said 2nd flotation concentrate.

  According to the present invention, the concentrate can be separated into a concentrate having a high copper concentration and a concentrate having a high molybdenum concentration.

It is a processing flow. It is an XRD analysis result of coarse beneficiation. It is a XRD analysis result of the coarse beneficiation after heat processing. It is a figure which shows Cu yield and Mo yield with respect to the flotation time in an Example. It is a figure which shows Cu yield and Mo yield with respect to the flotation time in a comparative example.

  Hereinafter, an embodiment for carrying out the present invention will be described.

(Embodiment)
The present embodiment includes a heating step of mixing crude sulfur obtained by pulverizing and sorting a chalcopyrite coarse ore containing molybdenum with simple sulfur and heating to 350 ° C. to 450 ° C., and the coarse obtained in the heating step. A separation step of crushing the concentrate and performing flotation, and separating the first flotation concentrate recovered in the previous stage of the flotation process and the second flotation concentrate recovered in the latter stage of the process And a method comprising: FIG. 1 is a processing flow.

(Crushing / sorting process)
First, 80% passing particle diameter of 100 to 300 μm is obtained at a copper grade of 3 to 15% by weight by crushing a chalcopyrite coarse ore containing molybdenum and having a copper grade of 0.3 to 3% by weight. Get the rough beneficiation. As the sorting method at this time, sieving, specific gravity sorting, table sorting, flotation or the like is adopted. In general, the coarse beneficiation includes silicate minerals such as quartz, pyrite (FeS ₂ ), copper sulfide ore, in addition to chalcopyrite and pyroxenite.

(Heating process)
Single sulfur is added to all or part of the coarse beneficiation and heated to 350 ° C. to 450 ° C. in an inert gas atmosphere to cause sulfidation, thereby obtaining a crude concentrate. By this sulfide conversion, chalcopyrite can be converted into copper indigo (CuS) and pyrite. The elemental sulfur is preferably added in an amount of 1.0 to 2.0 times the copper content of the coarse beneficiation target. If the amount of sulfur added during sulfidation is excessive, unreacted sulfur may cause a dust explosion when it is crushed or ground again. If the amount added is insufficient, the conversion of chalcopyrite will be insufficient. This is because the copper concentration in the molybdenite does not become low. By adding sulfur to chalcopyrite and heat treatment, chalcopyrite is converted to pyrite and copper indigo, but hydropyrite is heat-treated in an inert atmosphere and the mineral surface composition and shape change, leading to pyroxene lead. The floatability of the ore will change, and the yield and purity in the flotation process will be improved.

(Milling process)
By subjecting the coarse concentrate obtained in the heating step to crushing or grinding, the 80% passing particle diameter is adjusted to about 30 to 100 μm. By setting the particle size range of the coarse concentrate in this way, it is possible to improve the separation between the concentrate having a high copper concentration and the concentrate having a high molybdenum concentration during the flotation.

(Cu concentration process)
Next, the crude concentrate obtained in the milling process is subjected to flotation. In this flotation process, a process adopted in a normal copper mine is applied. That is, 100 to 200 g / t of a collecting agent is added, and air is blown into the bath liquid at pH 10.0 to 12.5. In flotation, the concentrate with high copper concentration floats preferentially, and then the concentrate with high molybdenum concentration floats. Then, the concentrate (copper concentrate) with high copper concentration can be collect | recovered by collect | recovering the floss (1st flotation concentrate) of the front | former stage of the process of flotation. Copper concentrate is subjected to a copper smelting process.

(Mo concentration process)
By recovering the floss (second flotation concentrate) in the latter stage of the flotation process, it is possible to recover the concentrate (molybdenum concentrate) having a high molybdenum concentration. Thereby, the crude concentrate can be separated into copper concentrate and molybdenum concentrate. In addition, 100 g / t or more of the collection agent is added again to the tank liquid (tailing slurry) after collecting the first flotation concentrate, and the flotation is continued with the tank liquid having a pH of 10.0 to 12.5. By doing so, the molybdenum concentrate can be efficiently recovered. By repeatedly subjecting this molybdenum concentrate to flotation, impurities such as copper can be further removed. When the copper grade of the obtained molybdenum concentrate exceeds 1% by weight, the copper content can be selectively leached by heat leaching with iron chloride or the like. The obtained molybdenum concentrate is subjected to dehydration and drying, and then subjected to a molybdenum smelting process.

  According to the present embodiment, by separating the first flotation concentrate collected in the first stage of the flotation process and the second flotation concentrate collected in the second stage of the flotation process, The ore can be separated into a concentrate with a high copper concentration and a concentrate with a high molybdenum concentration. In addition, it becomes possible to reduce the amount of slurry repeatedly processed by using the first-stage floss for copper smelting, the second-stage floss for molybdenum smelting, and the intermediate floss again for flotation.

  Hereinafter, a more specific description will be given with reference to examples. The present invention is not limited to these examples.

(Example)
Rough beading of chalcopyrite accompanied by molybdenite (Cu grade: 11.8 wt%, Mo grade: 0.26%, Fe grade: 21.4 wt%, sulfur grade: 24.4 wt%, passing 80% With respect to the particle diameter: 178 μm), simple sulfur was added by 1.0 times the weight of the coarsely processed copper, and a heating process was performed at 400 ° C. for 60 minutes in an N ₂ gas atmosphere. Crude concentrate after heating process (Cu grade: 10.9 wt%, Mo grade: 0.22%, Fe grade: 20.5 wt%, sulfur grade: 28.7 wt%, 80% passing particle size: 187 μm ) Shows that the pyrite has been converted into copper indigo and pyrite as shown in the analysis results by XRD (Rint-2200, manufactured by Rigaku Corporation) in FIGS. 2 represents the XRD analysis result of the coarse beneficiation, and FIG. 3 represents the analysis result after the heat treatment of the coarse beneficiation.

  Next, the coarse concentrate obtained in the heating step was milled with a wet ball mill so that the 80% passing particle diameter was 53 μm, and then a flotation process was performed. A Fahrenwald flotation machine was used as the flotation machine. The flotation reagent used was isopropyl xanthate (IPX) as a collection agent, methyl isobutyl carbinol (MIBC) as a foaming agent, and slaked lime as a pH adjuster.

  The procedure of the above flotation process is described below. Using slaked lime, the pH of the milled slurry was adjusted to 12.0, and 100 g / t of IPX was added to the crude concentrate, followed by conditioning for 3 min. After conditioning, 50 g / t of MIBC was added and air was introduced into the slurry. A froth layer was formed on the top of the slurry, and this was recovered. Floss was divided for every collection time and collected for 10 minutes. After recovery, IPX was added again at 50 g / t, and MIBC was added again at 20 g / t, and floss was recovered for 10 minutes. After recovery, IPX was added again at 100 g / t and MIBC at 20 g / t, and the floss was recovered for 10 minutes. The recovered floss and the slurry remaining in the flotation cell were filtered and dried, and weighed and subjected to elemental analysis. In elemental analysis, it was eluted after being melted together with sodium peroxide and sodium carbonate, diluted appropriately, and the concentration was measured with ICP-AES (Seiko Instrumental HVR1700) to determine the content.

(Comparative example)
In the comparative example, the coarse ore of chalcopyrite accompanied by hydropyrite (Cu grade: 10.1 wt%, Mo grade: 0.24%, Fe grade: 18.9 wt%, sulfur grade: 20.9 wt%) , 80% passing particle diameter: 173 μm) was milled with a wet ball mill to adjust the 80% passing particle diameter to 50 μm, and then the flotation process was carried out. A Fahrenwald flotation machine was used as the flotation machine. The flotation reagent used was methyl isobutyl carbinol (MIBC) as a foaming agent and slaked lime as a pH adjuster.

  The procedure of the above flotation process is described below. Using slaked lime, the pH of the milled slurry was adjusted to 11.0, and 12 g / t of MIBC was added to the crude concentrate, followed by conditioning for 1 min. After conditioning, air was introduced into the slurry. A froth layer was formed on the top of the slurry, and this was recovered. Floss was divided for every collection time and collected for 30 minutes. The recovered floss and the slurry remaining in the flotation cell were filtered and dried, and weighed and subjected to elemental analysis. In elemental analysis, it was eluted after being melted together with sodium peroxide and sodium carbonate, diluted appropriately, and the concentration was measured with ICP-AES (Seiko Instrumental HVR1700) to determine the content.

FIG. 4 shows Cu yield and Mo yield relative to flotation time in the examples. FIG. 5 shows the Cu yield and the Mo yield with respect to the flotation time in the comparative example. Table 1 shows concentrates of flotation times of 0 to 10, 10 to 20, and 20 to 30 minutes, Cu grades, Mo grades, Cu yields, and Mo yields of final tailings in Examples. Table 2 shows concentrates of flotation times 0 to 10, 10 to 20, and 20 to 30 min in comparative examples, Cu quality, Mo quality, Cu yield, and Mo yield of the final tailings.

  In the comparative example, as can be seen from FIG. 5, the Cu yield and the Mo yield are similar to the flotation time. Therefore, the crude concentrate could not be separated into a concentrate having a high copper concentration and a concentrate having a high molybdenum concentration. Also from Table 2, concentrates with high Mo grade (Comparative Examples 0 to 10 min) also had high Cu grade.

  On the other hand, in the examples, as can be seen from FIG. 4, the Cu yield was high in the flotation time 0 to 20 min (collecting agent addition amount 100 to 150 g / t), but the Mo yield was low. The Mo yield was increased in the flotation time of 20 min to 30 min (re-addition of collection agent 100 g / t). From Table 1, copper concentrates with low Mo quality could be recovered in concentrates with a flotation time of 0 to 20 min (collecting agent addition amount of 100 to 150 g / t). Further, the concentrate with the flotation time of 20 to 30 min (re-addition of the collection agent 100 g / t) had the highest Mo grade and the Cu grade was as low as 5.8%. Therefore, molybdenum concentrate with low Cu quality could be recovered. By adding sulfur to chalcopyrite and heat treatment, chalcopyrite is converted to pyrite and copper indigo, but hydropyrite is heat-treated in an inert atmosphere and the mineral surface composition and shape change, leading to pyroxene lead. It is thought that the floating property of the ore changed and the yield and purity in the flotation process were improved. The molybdenum concentrate may be subjected to flotation again to remove impurities such as copper and concentrate the hydropyrite. Alternatively, copper can be selectively leached by heat leaching with iron chloride or the like.

  Although the embodiments of the present invention have been described in detail above, the present invention is not limited to such specific embodiments, and various modifications and changes can be made within the scope of the gist of the present invention described in the claims. It can be changed.

Claims (8)

A heating step in which a coarse bead obtained by pulverizing and selecting a chalcopyrite coarse ore containing molybdenum is mixed with simple sulfur and heated to 350 ° C. to 450 ° C .;
The coarse concentrate obtained in the heating step is crushed and subjected to flotation, and the first flotation concentrate recovered in the previous stage of the flotation process and the second flotation recovered in the latter stage of the process. And a separation step of separating the concentrate. Subjecting the first flotation concentrate to a copper smelting process;
The separation method according to claim 1, wherein the second flotation concentrate is subjected to a molybdenum smelting step.   The separation method according to claim 1 or 2, wherein the chalcopyrite coarse ore contains pyrite.   The said chalcopyrite coarse ore contains at least 1 sort (s) among pyrite, a chalcopyrite, and copper indigo, The separation method as described in any one of Claims 1-3 characterized by the above-mentioned.   The separation method according to any one of claims 1 to 4, wherein the coarse beneficiation has a copper grade of 3 to 15% by weight and an 80% passing particle size of 100 to 300 µm.   The said heating process WHEREIN: 1.0 weight times-2.0 weight times single-piece | unit sulfur are added with respect to the copper content in the said rough beneficiation, The 1 item | term as described in any one of Claims 1-5 characterized by the above-mentioned. Separation method.   The said separation process WHEREIN: After crushing the said rough concentrate until an 80% passage particle diameter becomes 30-100 micrometers, the said flotation is performed, It is characterized by the above-mentioned. Separation method.   In the separation step, after recovering the first floating concentrate, the collection material is added to the floating beneficiation tank in an amount of 100 g / t or more, and the pH of the tank liquid is adjusted to 10.0 to 12.5. It collect | recovers as said 2nd flotation concentrate, The separation method as described in any one of Claims 1-7 characterized by the above-mentioned.

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