AU2004200260A1 - Method of Recovering Sulfur from Leaching Residue under Atmospheric Pressure - Google Patents
Method of Recovering Sulfur from Leaching Residue under Atmospheric Pressure Download PDFInfo
- Publication number
- AU2004200260A1 AU2004200260A1 AU2004200260A AU2004200260A AU2004200260A1 AU 2004200260 A1 AU2004200260 A1 AU 2004200260A1 AU 2004200260 A AU2004200260 A AU 2004200260A AU 2004200260 A AU2004200260 A AU 2004200260A AU 2004200260 A1 AU2004200260 A1 AU 2004200260A1
- Authority
- AU
- Australia
- Prior art keywords
- sulfur
- leaching
- flotation
- residue
- leaching residue
- 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.)
- Granted
Links
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 title claims description 92
- 229910052717 sulfur Inorganic materials 0.000 title claims description 92
- 239000011593 sulfur Substances 0.000 title claims description 90
- 238000000034 method Methods 0.000 title claims description 75
- 238000002386 leaching Methods 0.000 title claims description 58
- 238000005188 flotation Methods 0.000 claims description 83
- 230000008569 process Effects 0.000 claims description 44
- 239000010949 copper Substances 0.000 claims description 42
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 40
- 229910052802 copper Inorganic materials 0.000 claims description 36
- 238000012545 processing Methods 0.000 claims description 32
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims description 26
- 239000012141 concentrate Substances 0.000 claims description 20
- 239000002253 acid Substances 0.000 claims description 17
- 230000005484 gravity Effects 0.000 claims description 13
- 229910052500 inorganic mineral Inorganic materials 0.000 claims description 10
- 239000011707 mineral Substances 0.000 claims description 10
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 claims description 9
- 238000000926 separation method Methods 0.000 claims description 7
- 229910052751 metal Inorganic materials 0.000 claims description 3
- 238000005065 mining Methods 0.000 claims description 2
- 150000003568 thioethers Chemical class 0.000 claims 2
- 239000000243 solution Substances 0.000 description 32
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 26
- 238000011084 recovery Methods 0.000 description 25
- 229910052742 iron Inorganic materials 0.000 description 16
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 13
- 150000004763 sulfides Chemical class 0.000 description 12
- 239000002245 particle Substances 0.000 description 8
- 229910002588 FeOOH Inorganic materials 0.000 description 6
- GRYLNZFGIOXLOG-UHFFFAOYSA-N Nitric acid Chemical compound O[N+]([O-])=O GRYLNZFGIOXLOG-UHFFFAOYSA-N 0.000 description 6
- 229910017604 nitric acid Inorganic materials 0.000 description 6
- 238000007670 refining Methods 0.000 description 6
- WVYWICLMDOOCFB-UHFFFAOYSA-N 4-methyl-2-pentanol Chemical compound CC(C)CC(C)O WVYWICLMDOOCFB-UHFFFAOYSA-N 0.000 description 5
- 238000010306 acid treatment Methods 0.000 description 5
- 238000012216 screening Methods 0.000 description 5
- VEXZGXHMUGYJMC-UHFFFAOYSA-M Chloride anion Chemical compound [Cl-] VEXZGXHMUGYJMC-UHFFFAOYSA-M 0.000 description 4
- DVRDHUBQLOKMHZ-UHFFFAOYSA-N chalcopyrite Chemical compound [S-2].[S-2].[Fe+2].[Cu+2] DVRDHUBQLOKMHZ-UHFFFAOYSA-N 0.000 description 4
- 229910052951 chalcopyrite Inorganic materials 0.000 description 4
- 238000009854 hydrometallurgy Methods 0.000 description 4
- 239000007788 liquid Substances 0.000 description 4
- 238000002441 X-ray diffraction Methods 0.000 description 3
- 229910001779 copper mineral Inorganic materials 0.000 description 3
- 239000010419 fine particle Substances 0.000 description 3
- 238000004519 manufacturing process Methods 0.000 description 3
- 239000002904 solvent Substances 0.000 description 3
- 238000007664 blowing Methods 0.000 description 2
- 230000003247 decreasing effect Effects 0.000 description 2
- 238000005868 electrolysis reaction Methods 0.000 description 2
- 238000007667 floating Methods 0.000 description 2
- 230000002209 hydrophobic effect Effects 0.000 description 2
- 239000008235 industrial water Substances 0.000 description 2
- 238000009853 pyrometallurgy Methods 0.000 description 2
- 238000010792 warming Methods 0.000 description 2
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 2
- 229910014265 BrCl Inorganic materials 0.000 description 1
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 description 1
- 230000009471 action Effects 0.000 description 1
- 238000013459 approach Methods 0.000 description 1
- CODNYICXDISAEA-UHFFFAOYSA-N bromine monochloride Chemical compound BrCl CODNYICXDISAEA-UHFFFAOYSA-N 0.000 description 1
- 238000006243 chemical reaction Methods 0.000 description 1
- 239000003795 chemical substances by application Substances 0.000 description 1
- 150000003841 chloride salts Chemical class 0.000 description 1
- 238000007796 conventional method Methods 0.000 description 1
- 238000009826 distribution Methods 0.000 description 1
- 230000005611 electricity Effects 0.000 description 1
- 230000007613 environmental effect Effects 0.000 description 1
- 238000009776 industrial production Methods 0.000 description 1
- 239000000463 material Substances 0.000 description 1
- 238000002844 melting Methods 0.000 description 1
- 230000008018 melting Effects 0.000 description 1
- 239000002184 metal Substances 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 230000009467 reduction Effects 0.000 description 1
- 238000011160 research Methods 0.000 description 1
- 239000012266 salt solution Substances 0.000 description 1
- 239000011701 zinc Substances 0.000 description 1
- 229910052725 zinc Inorganic materials 0.000 description 1
Classifications
-
- 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
Landscapes
- Manufacture And Refinement Of Metals (AREA)
- Separation Of Solids By Using Liquids Or Pneumatic Power (AREA)
Description
S&F Ref: 663098
AUSTRALIA
PATENTS ACT 1990 COMPLETE SPECIFICATION FOR A STANDARD PATENT Name and Address of Applicant: Actual Inventor(s): Address for Service: Invention Title: Nippon Mining Metals Co., Ltd.
10-1, Toranomon 2-chome, Minato-ku Tokyo 105-0001 Japan Atsushi Saito Spruson Ferguson St Martins Tower Level 31 Market Street Sydney NSW 2000 (CCN 3710000177) Method of Recovering Sulfur from Leaching Residue under Atmospheric Pressure The following statement is a full description of this invention, including the best method of performing it known to me/us:- 5845c TITLE OF THE INVENTIQN Method of Recovering Sulfur from Leaching Residue under Atmospheric Pressure BACKGROUND OF THE INVENTION 1. Field of the invention The present invention relates to a method of recovering sulfur from a residue that is generated by leaching sulphide minerals copper concentrates) under atmospheric pressure.
2. Description of the Related Art In earlier times, the pyrometallurgical process had been predominantly used as a copper refining process. However, recently, an SX-EW process, which is one of the hydrometallurgical processes, has become remarkably popular as the copper refining process to extract copper typically from cooper oxide ores or secondary copper ores, and the process has achieved about 20% of the world copper production.
When producing copper metal at copper mine site, they prefer the hydrometallurgical process to the pyrometallurgical process because of its lower production cost and an easier environmental protection action. This is the reason why the hydrometallurgical process has made progress in producing copper. However, the hydrometallurgical process has not yet reached a technical level to process main copper mineral such as primary copper mineral and chalcopyrite. Several researches have been carried out to achieve a technical breakthrough of processing the primary copper mineral and the chalcopyrite.
One of the hydrometallurgical copper refining processes, such as an Intec copper process using a chloride salt solution, is a promising method of refining the primary copper material, since the method can facilitate a leaching operation and reduce electricity consumption of univalent copper 1 electrolysis by half. Therefore, there is a need of efficiently recovering sulfur from a chloride leaching residue generated by the hydrometallurgical copper refining process.
For example, a leaching process used in the Intec copper process leaches the chalcopyrite from Cu(2+) and BrCl 2 contained in a solution after electrolysis, and produces a leaching liquid containing and a leaching residue in the form of FeOOH and elemental sulfur. Therefore, there is also a need of separating sulfur and iron from the leaching residue in an industrially and economically viable manner.
Another known the hydrometallurgical copper refining process has acid leaching where the copper concentrate is leached using Fe(3+) at room temperature and under atmospheric pressure. However, this acid leaching also has a problem of recovering sulfur contained in the leaching residue.
Conventional methods of separating and recovering sulfur from the residue can be classified into two categories: a first method having the steps of extracting sulfur using a solvent and separating sulfur from iron, and a second method including pressurization and screening (flotation) The first method that uses the solvent is not preferable to recover sulfur, because the solvent is costly and not easy to handle in an industrially simple manner. The second method that includes the pressurization is not feasible because of a high cost of the pressurization. Consequently, both methods are not industrially and economically viable.
Japanese Patent Publication No. 6-43619, "PROCESS FOR THE LEACHING OF SULPHIDES CONTAINING ZINC AND IRON", discloses a leaching process based on a combination of leaching under pressure and leaching under atmospheric pressure. It is to be noted that the leaching process of the cited patent necessitates the leaching under 2 pressure applied.
It is known that, in case of leaching of sulphide minerals copper concentrates) under pressure, sulfur that remains in a residue is easily recovered from the residue by flotation. Thus, the leaching process of the cited patent involves the leaching under pressure. At the time of the leaching under pressure, the sulfur contained in the residue may easily be recovered by the flotation. A pressurization process causes sulfur to reach its melting point. Once the sulfur has fused, the sulfur can re-crystallize in the form of a certain size of particles and rather suitable to float in the liquid.
On the contrary, since the sulfur contained in the residue after leaching under atmospheric pressure remains in the form of fine particles, it is not easy to selectively separate the sulfur by the flotation.
SUMMARY OF THE INVENTION It is an object of the present invention to provide a method of economically and industrially recovering sulfur from a leaching residue produced by leaching sulphide minerals copper concentrates) under atmospheric pressure, in which the above disadvantages are eliminated.
The inventor of the present invention has studied various methods of recovering sulfur from a residue and has reached the following invention.
The object of the present invention is achieved by a method of recovering sulfur from a leaching residue produced by leaching sulphide minerals (e.g.
copper concentrates) under atmospheric pressure, the method comprising the steps of: pretreating the leaching residue with an acid solution; and selectively recovering sulfur from the leaching residue by a flotation process.
3 According to an embodiment of the present invention, the acid solution is a sulfuric acid solution.
According to an embodiment of the present invention, the step of pretreating the leaching residue with an acid solution comprises the step of: pretreating the leaching residue with a treatment solution of sulfuric acid having a pH value equal to or less than 1.5, at a pulp density of 10 to 50 mass percent, for a processing time of 20 to 60 minutes, and at a processing temperature of 30 to 90 oC.
According to an embodiment of the present invention, the step of selectively recovering sulfur from the leaching residue by a flotation process comprises the step of: applying the Rougher flotation to the pretreated residue at a pulp density of 10 to 35 mass percent, in a treatment solution having a pH value equal to or less than 2.2, and at a processing temperature of 20 to 90 °C.
According to a further embodiment of the present invention, the step of selectively recovering sulfur from the leaching residue by a flotation process further comprises the step of: applying the cleaner flotation to a product produced by the floatation process.
According to a still further embodiment of the present invention, the step of selectively recovering sulfur from the leaching residue by a flotation process further comprises the step of: if a product containing sulfur produced by the cleaner flotation includes sulfides which have not leached away, applying a gravity separation process to the product so as to selectively removing the sulfides from the product.
4 BRIEF DESCRIPTION OF THE DRAWINGS Other objects, features and advantages of the present invention will become more apparent from the following detailed description when read in conjunction with the accompanying drawings, in which: Fig. 1 shows results of an x-ray diffraction analysis (XRD) according to the present invention; Fig. 2 shows a particle size distribution of feed (Cu concentrate) and residue from a processing ore according to the present invention; Fig. 3 plots a sulfur recovery percentage in relation to a warming time during a pretreatment according to the present invention; Fig. 4 plots a sulfur recovery percentage in relation to a pH during the pretreatment according to the present invention; Fig. 5 plots a sulfur recovery percentage in relation to a warming temperature during the pretreatment according to the present invention; Fig. 6 plots a sulfur recovery percentage in relation to a pH during rougher flotation; Fig. 7 plots a sulfur recovery percentage in relation to sulfur grade at each process of the present invention; and Fig. 8 shows a processing flow according to one embodiment of the present invention.
DESCRIPTION OF THE PREFERRED EMBODIMENTS The present invention will be described in detail, by way of an example, with reference to the accompanying drawings. It should be noted that the present invention is not limited to the specifically disclosed embodiments.
A processing object of the present invention, that is to say a leaching residue, is composed of, for example, 10 to 35 mass% Fe (iron), 15 to 35 mass% S (sulfur) and 0.7 to 1.2 mass% Cu (copper).
5 If the residue is a chloride leaching residue, then Fe is leached out of chalcopyrite and exists in the form of FeOOH, and most of S exists in the form of elemental sulfur, as can be seen from the results of an x-ray diffraction analysis shown in Fig. 1. The residue contains some ores which have not yet leached away.
Sulfur essentially has a native hydrophobic property, whereas FeOOH is very hydrophilic. Therefore, if particles of sulfur and FeOOH are physically separated from each other, it is appreciated that sulfur and FeOOH can be separated from each other by flotation utilizing a difference between hydrophilic and hydrophobic aspects.
With reference to Fig. 2, however, the particle size of the residue is much smaller than that of copper concentrates prior to a leaching process, and has a very small 50% passing size of between 2 and 4 microns.
The floatation process is not capable of separating fine particles which are smaller than an applicable particle size limit, because such fine particles will not attach to an air bubble. The particle size limit is said to be between 20 and microns for ordinary ores and 7 microns for MacArthur River mineral ores which probably has the smallest grain size limit.
For the reasons stated above, though the flotation process can be performed at attractive costs, it is a very difficult approach to reduction in practice in technical point of view.
According to one aspect of the present invention, the residue is previously pretreated with an acid solution. Preferably, the acid solution is a sulfuric acid solution, as described later.
Preferably, a ratio of the leaching residue to the sulfuric acid solution is such that a pulp density is equal to 10 to 35 mass%. This is because industrial 6 production efficiency is low at the pulp density below mass% and the pretreatment process becomes difficult to carry out above the pulp density of 35 mass% as the solution will not be sufficiently stirred.
The time of a pretreatment process is preferably set to be between 20 and 60 minutes, as shown in Fig. 3.
Otherwise, a final sulfur recovery percentage will not go beyond 90 mass%, because the pretreatment with the sulfuric acid solution is insufficient at the processing time below 20 minutes and the industrial production efficiency becomes inferior at the processing time over 60 minutes.
More preferably, the processing time is set to be between 30 and 60 minutes. This is because the sulfur recovery percentage during a first flotation process (hereinafter referred to as "rougher flotation") amounts to 95 mass% or more.
Properties of several acid treatment solutions, such as sulfuric acid, hydrochloric acid and nitric acid, used in the pretreatment process are shown in Table I. Under a treating condition where an identical pH value is used, the sulfur recovery percentages for the sulfuric acid, hydrochloric acid and nitric acid are 96.0 mass%, 44.8 mass% and 58.5 mass%, respectively.
An addition of the hydrochloric acid or the nitric acid improves the sulfur recovery percentage in comparison with the case where no acid is added. Preferably, the sulfuric acid solution is used as the acid solution in the pretreatment process in terms of its sulfur recovery percentage. Furthermore, the sulfuric acid is more cost effective and easier to handle in later processes than the hydrochloric acid and the nitric acid.
Of course, the sulfur recovery percentage may be improved with increasing acid concentration of the hydrochloric acid or the nitric acid.
It can be seen from the property of the solution 7 after the pretreatment process that the solution produced using the sulfuric acid has a higher iron concentration than using the hydrochloric acid or the nitric acid. From this point of view, we can consider that FeOOH covering a surface of sulfur particle elutes by a reaction represented by the following formula (1) and is removed from the surface of the sulfur particle, so that an essential floating characteristic of sulfur is revealed and improved.
2FeOOH+3H 2
SO
4 -Fe 2
(SO
4 3 +4H 2 0 (1) Table I Type Kinds Grade (mass%) Recovery (mass%) of of S Fe Cu Wt S Fe Cu Acid Ores No F 27.9 30.3 1.1 100.0 100.0 100.0 100.0 Acid C 47.2 30.4 1.1 22.2 37.7 22.3 22.7 T 22.3 30.2 1.1 77.8 62.3 77.7 77.3
H
2
SO
4 F 30.0 29.0 1.0 100.0 100.0 100.0 100.0 C 52.4 21.6 1.5 55.0 96.0 41.0 83.0 T 2.7 38.0 0.4 45.0 4.0 59.0 17.0
HNO
3 F 29.9 29.2 0.9 100.0 100.0 100.0 100.0 C 53.3 26.4 1.3 32.8 58.5 29.7 47.9 T 18.5 30.5 0.7 67.2 41.5 70.3 52.1 HCl F 29.9 29.9 0.9 100.0 100.0 100.0 100.0 C 60.2 23.9 1.5 22.2 44.8 17.8 34.1 T 21.2 31.6 0.8 77.8 55.2 82.2 65.9 indicates indicates indicates feed (residue) concentrate.
tailing.
Preferably, the sulfuric acid treatment solution for the pretreatment is adjusted such that its pH value is equal to or less than 1.5, as shown in Fig. 4. This is because the sulfur recovery percentage reaches mass% or more during rougher flotation. More 8 preferably, the pH value of the sulfuric acid treatment solution is adjusted to be less than 1.0, because the sulfur recovery percentage reaches 95 mass% or more during rougher flotation.
The processing temperature during the pretreatment is preferably between 30 and 90 'C, because the sulfur recovery percentage reaches 90 mass% or more during rougher flotation. More preferably, the processing temperature during the pretreatment is between 40 and 90 oC, because the sulfur recovery percentage reaches 95 mass% or more during rougher flotation at a processing temperature higher than 40 'C.
After the pretreatment process, the pulp density during rougher flotation is preferably between 10 and 35 mass%. This is because the rougher flotation is not industrially efficient at the pulp density below mass% and the viscosity of the solution is too high to facilitate the rougher flotation at the pulp density above 35 mass%.
The processing time for the pretreatment with the sulfuric acid solution is preferably between 5 and minutes. This is because the sulfur recovery percentage will not reach 90 mass% due to the insufficient pretreatment with the sulfuric acid solution at the processing time below 5 minutes, and the industrial efficiency is inferior at the processing time above 75 minutes.
More preferably, the processing time is between and 75 minutes, because the sulfur recovery percentage reaches 95 mass% or more at the processing time above 30 minutes during rougher flotation.
Preferably, the sulfuric acid treatment solution for rougher flotation is adjusted such that its pH value is equal to or less than 2.2, as shown in Fig. 6.
This is because the sulfur recovery percentage reaches mass% or more during rougher flotation. More preferably, the pH value of the sulfuric acid treatment 9
I
solution is adjusted to be between 1.5 and 2.1, because the sulfur recovery percentage reaches 95 mass% or more during rougher flotation.
The processing temperature during rougher flotation is preferably between 20 and 90 0 C, because the sulfur recovery percentage reaches 90 mass% or more during rougher flotation. More preferably, the processing temperature during coarse screening is between 40 and 90 because the sulfur recovery percentage reaches 95 mass% or more during coarse screening at the processing temperature above 40 'C.
It is noted that the processing temperature above 90 0 C will not be effective to further improve the sulfur recovery percentage.
It is obvious that a frother, such as Methyl Isobutyl Carbinol (MIBC), of 40 to 80 g/t (g/ton) is added to the residue in dried form.
In the rougher flotation, a floatation machine, such as a column flotation machine, may be used.
After the rougher flotation, the cleaner flotation may be performed. Preferably, the, cleaner flotation repeats cleaner flotation two or more times.
This is because grade of the recovered sulfur reaches about 80 mass% or more, as shown in Fig. 7.
Since forming property of the forther added to the dry residue in the rougher flotation is maintained during cleaner flotation, there is no need of further adding the forther to the residue. In stead of adding the agent, air bubbles are kept in a floating position by applying air blowing of 250 to 700 1/min/m 2 to the residue and cleaner flotation is applied to the residue in normal industrial water.
The column floatation machine may also be used as the flotation machine for cleaner flotation.
Grade of a product containing sulfur generated by cleaner flotation is as follow: FIRST FINE CLEANER FLOTATION: S grade is 65 to 69 10 mass%, Fe grade is 13.0 to 17.0 mass%, and Cu grade is between 1.7 to 2.1 mass%; and SECOND FINE CLEANER FLOTATION: S grade is 77 to 81 mass%, Fe grade is 3.0 to 7.0 mass%, and Cu grade is 1.6 to 1.9 mass%. It can be seen that S grade at the second stage is higher than that of the first stage.
If the Fe grade and/or the Cu grade of the product containing sulfur are high, gravity flotation, or gravity separation may be performed in order to recover sulfides from the product, as the sulfides have not yet leached away. Since the specific gravity of sulfides is about 5 and the specific gravity of sulfur is about 2, the sulfides are subject to the gravity flotation in water using a plate inclined at an angle of about 0.5 to 1.0 degrees and by applying a vibration having the number of vibrations of about 230 to 280 rpm and vibration width of 13 to 25 mm.
Alternatively, the gravity flotation may be performed by a gravity flotation machine based on centrifugal separation.
After this gravity flotation, on one hand, the grade of a resultant sulfur concentrate is such that S grade is 78.0 to 82.0 mass%, Fe grade is 3.0 to 3.8 mass%, and Cu grade is 1.0 to 1.4 mass%. On the other hand, the grade of the sulfides which have not yet leached away is such that S grade is 40.0 to 48.0 mass%, Fe grade is 34.0 to 38.0 mass%, and Cu grade is 13.0 to 17.0 mass%. It can be seen that sulfur grade of the sulfur concentrate is increased and Cu grade and Fe grade are decreased.
As will be apparent from the above mentioned description, the present invention provides a method of recovering sulfur from a leaching residue produced by leaching sulphide minerals copper concentrates) simply under atmospheric pressure.
EXAMPLE
11 Referring to Figs. 7 and 8, an embodiment of the present invention will be explained, by way of an example, in conjunction with Table II representing processing objects and processing results for each step of a method of recovering sulfur from a leaching residue according to the embodiment of the present invention.
A processing object, that is to say, a chloride leaching residue of a copper concentrate contains S, Fe and Cu such that S grade is 30.0 mass%, Fe grade is 29.0 mass%, and Cu grade is 1.0 mass%, as shown in Table II.
Table II Object Grade (mass%) Recovery (mass%) S Fe Cu Wt S Fe Cu 30.0 29.0 1.0 100.0 100.0 100.0 100.0 54.0 22.0 1.6 53.0 95.4 40.2 84.8 67.0 15.0 1.9 42.0 93.8 21.7 79.8 79.0 5.0 1.8 33.8 89.0 5.8 60.8 80.6 3.4 1.2 32.3 86.8 3.8 38.4 44.5 38.6 15.0 1.5 2.2 2.0 22.4 Object indicates a residue or a processing object.
Object indicates a concentrate produced by rougher flotation.
Object indicates a product produced by the first stage of cleaner flotation.
Object indicates a product produced by the second stage of cleaner flotation.
Object indicates a resultant sulfur concentrate.
Object indicates sulfides which have not leached away.
The processing object was pretreated with a sulfuric acid solution under treatment conditions: 12 pH=l.0, a pulp density of 30 mass%, a treatment liquid temperature of 80 and a processing time of minutes.
Then, the object was subject to rougher flotation under rougher flotation conditions: pulp density of 30 mass%, treatment liquid temperature of 20 processing time (flotation time) of minutes.
Methyl Isobutyl Carbinol (MIBC) was used as a frother. 60g/t of MIBC was added to the dry residue.
This rougher flotation was carried out in a column flotation machine.
The resultant object after rougher flotation was the rougher flotation concentrate such that S grade was 54 mass%, Fe grade was 22.0 mass%, and Cu grade was 1.6 mass%, as shown in the object of Table II and Fig. 7.
It can be seen that this rougher flotation increased S grade of the object to about twice that of the object Subsequently, the cleaner flotation was applied to the object and thus producing the object as shown in Fig. 8.
During fine screening, since air bubbles generated by the previous rougher flotation were flowed into a solution, industrial water was used as the solution and air blowing of 318 i/min/m 2 0.1 l/min/3.14cm 2 was introduced from a lower part of a flotation machine in order to move the air bubbles upwards.
The solution was kept at a room temperature of
°C.
A column flotation machine was used as the flotation machine.
The fine screening process included two stages of cleaner flotation. After the first stage, the resultant object was composed as follows: S grade 13 of 67.0 mass%, Fe grade of 15.0 mass%, and Cu grade of 1.9 mass%, as shown in Table II and Fig. 7. Sulfur grade of the object was more increased than that of the object produced by rougher flotation.
After the second stage of the cleaner flotation process, the resultant object was composed as follows: S grade of 79.0 mass%, Fe grade of 5.0 mass%, and Cu grade of 1.8 mass%, as shown in Table II and Fig.
7. Sulfur grade of the object was more increased than that of the object produced by rougher flotation and iron grade of the object was lower than that of the object (3) In addition, gravity separation using the shaking table inclined at an angle of about 0.5 degrees was applied to a product containing sulfur, i.e. the object generated by the second cleaner flotation stage, with vibration of 250 rpm and vibration width of 20 mm in water.
In this manner, sulfides that have not leached away and a sulfur concentrate were produced, as shown in objects and of Table II. On one hand, the sulfides contained Cu such that Cu grade is 15 mass%.
On the other hand, the sulfur concentrate contained S such that S grade much increased to 80.6 mass% and F grade and Cu grade decreased, and thus resulted in the desirable sulfur concentrate.
As apparent from steps to of Fig. 7, sulfur grade increased as the step. proceeded. It is noted that numbers to of Fig. 7 correspond to those of Fig. 8. The object numbers to in Table II also correspond to the numbers in Fig. 8.
Treatment conditions for each of the steps should be determined so that desired sulfur recovery percentage and sulfur grade can be achieved.
In the above mentioned example, the method of recovering sulfur from a leaching residue according to one embodiment of the present invention includes a 14 cleaner flotation having two stages and an additional gravity separation process. However, the present invention is not limited to the specifically disclosed embodiments, and variations and modifications may be made without departing from the scope of the present invention mentioned before.
According to an embodiment of the present invention, since a leaching residue that has been leached out of sulphide minerals copper concentrates) is pretreated with an acid solution, high-grade sulfur can be recovered from a chloride leaching residue using a simple flotation process.
According to an embodiment of the present invention, since a sulfuric acid solution is used as the acid solution, sulfur grade is further improved.
According to a further embodiment of the present invention, since a flotation process includes rougher flotation and cleaner flotation to separate sulfur from the residue, a product containing sulfur with a high sulfur concentration can be recovered from the residue.
According to a still further embodiment of the present invention, since the method further comprises a gravity separation process after the flotation process, sulfides that have not leached away can be efficiently removed from the product and more high-grade sulfur can be achieved.
15
Claims (8)
1. A method of recovering sulfur from a leaching residue produced by leaching sulphide minerals under atmospheric pressure, the method comprising the steps of: pretreating the leaching residue with an acid solution; and selectively recovering sulfur from the leaching residue by a flotation process.
2. The method of claim 1, wherein said sulphide minerals are copper concentrates.
3. The method of claim 1 or claim 2, wherein the acid solution is a sulfuric acid solution.
4. The method of claim 1 or claim 2, wherein the step of pretreating the leaching 1o residue with an acid solution comprises the step of: pretreating the leaching residue with a treatment solution of sulfuric acid having a pH value equal to or less than 1.5, at a pulp density of 10 to 50 mass percent, for a processing time of 20 to 60 minutes, and at a processing temperature of 30 to 90 C. The method of any one of claims 1 to 4, wherein the step of selectively recovering sulfur from the leaching residue by a flotation process comprises the step of: applying the flotation process to the pretreated residue at a pulp density of 10 to mass percent, in a treatment solution having a pH value equal to or less than 2.2, and at a processing temperature of 20 to 90 'C.
6. The method of any one of claims 1 to 5, wherein the step of selectively recovering sulfur from the leaching residue by a flotation process further comprises the step of: applying a cleaner flotation to a product produced by the flotation process.
7. The method as claimed in claim 6, wherein the step of selectively recovering sulfur from the leaching residue by a flotation process further comprises the step of: if a product containing sulfur produced by the cleaner flotation includes sulfides which have not leached away, applying a gravity separation process to the product so as to selectively removing the sulfides from the product.
8. A method of recovering sulfur from a leaching residue produced by leaching sulphide minerals under atmospheric pressure, the method being substantially as hereinbefore described with reference to the Example.
9. Sulfur recovered from a leaching residue by a method of any one of claims 1 to 8. Dated 19 January, 2004 Nippon Mining Metals Co., Ltd. Patent Attorneys for the Applicant/Nominated Person SPRUSON FERGUSON A663098speci
Applications Claiming Priority (2)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| JP2003-050397 | 2003-02-27 | ||
| JP2003050397A JP4112401B2 (en) | 2003-02-27 | 2003-02-27 | Recovery method of sulfur from leaching residue by atmospheric pressure method |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| AU2004200260A1 true AU2004200260A1 (en) | 2004-09-16 |
| AU2004200260B2 AU2004200260B2 (en) | 2006-02-23 |
Family
ID=32905656
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| AU2004200260A Expired AU2004200260B2 (en) | 2003-02-27 | 2004-01-23 | Method of Recovering Sulfur from Leaching Residue under Atmospheric Pressure |
Country Status (3)
| Country | Link |
|---|---|
| JP (1) | JP4112401B2 (en) |
| AU (1) | AU2004200260B2 (en) |
| CA (1) | CA2454854C (en) |
Families Citing this family (10)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| JP5497723B2 (en) * | 2011-10-21 | 2014-05-21 | Ｊｘ日鉱日石金属株式会社 | Copper concentrate processing method |
| CN102513204A (en) * | 2011-12-21 | 2012-06-27 | 大冶有色设计研究院有限公司 | Beneficiation method of sieving and flotation combination process for recycled copper of copper smelting converter slag |
| CN103894281B (en) * | 2012-12-28 | 2016-06-15 | 北京有色金属研究总院 | A kind of selecting smelting combination technique processing copper sulfide zinc and zinc oxide composite ore |
| CN103611624B (en) * | 2013-11-27 | 2015-11-25 | 中南大学 | A kind of flotation-acidleach process integration processing low-grade mixed copper ore |
| RU2544329C1 (en) * | 2014-02-14 | 2015-03-20 | Открытое акционерное общество "Горно-металлургическая компания "Норильский никель" | Processing method of pulp after autoclave-oxidising leaching of sulphide polymetallic materials, which contains iron oxides and elemental sulphur |
| CN105327771B (en) * | 2015-12-04 | 2017-06-23 | 云南锡业股份有限公司卡房分公司 | A kind of fine grinding and comprehensive reutilization ore-dressing technique method containing copper sulfide concentrate |
| CN109261345B (en) * | 2018-08-01 | 2021-10-22 | 昆明理工大学 | Copper-sulfur ore separation method |
| CN109954589A (en) * | 2019-04-23 | 2019-07-02 | 中南大学 | A method of the flotation recovery iron concentrate from high-alkali Pb-Zn tailings |
| CN111632744A (en) * | 2020-04-28 | 2020-09-08 | 西北矿冶研究院 | Beneficiation method for recovering copper sulfide from copper oxide acid leaching residues |
| CN115041303A (en) * | 2022-05-23 | 2022-09-13 | 中南大学 | Method for inhibiting lead and floating sulfur and strengthening flotation of high-sulfur slag |
Family Cites Families (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| GB1509537A (en) * | 1974-09-13 | 1978-05-04 | Cominco Ltd | Treatment of zinc plant residues |
| IT1231332B (en) * | 1989-07-31 | 1991-11-28 | Engitec Impianti | ELECTROLYTIC LEAD AND ELEMENTAL SULFUR PRODUCTION PROCESS FROM GALENA. |
-
2003
- 2003-02-27 JP JP2003050397A patent/JP4112401B2/en not_active Expired - Fee Related
- 2003-12-30 CA CA 2454854 patent/CA2454854C/en not_active Expired - Lifetime
-
2004
- 2004-01-23 AU AU2004200260A patent/AU2004200260B2/en not_active Expired
Also Published As
| Publication number | Publication date |
|---|---|
| JP4112401B2 (en) | 2008-07-02 |
| CA2454854A1 (en) | 2004-08-27 |
| CA2454854C (en) | 2012-08-07 |
| AU2004200260B2 (en) | 2006-02-23 |
| JP2004256363A (en) | 2004-09-16 |
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| JP3705815B2 (en) | Mineral leaching process at atmospheric pressure | |
| CA2683506C (en) | Process for precious metal recovery from a sulphide ore or concentrate or other feed material | |
| CN100485059C (en) | Method for recovering gold from sulphide concentrate | |
| AU670621B2 (en) | A method for recovering gold and other precious metals from carbonaceous ores | |
| US8052774B2 (en) | Method for concentration of gold in copper sulfide minerals | |
| US5792235A (en) | Method for recovering gold and other precious metals from carbonaceous ores | |
| JP2007297717A (en) | Method for recovering copper from sulfide ore material using high temperature pressure leaching, solvent extraction and electrowinning | |
| KR101399953B1 (en) | Method for producing copper concentrates from complex copper ore | |
| AU2004200260B2 (en) | Method of Recovering Sulfur from Leaching Residue under Atmospheric Pressure | |
| Agorhom et al. | Challenges and opportunities in the recovery/rejection of trace elements in copper flotation-a review | |
| CA2480229C (en) | Process for extracting platinum group metals | |
| AU752203B2 (en) | Method for recovering gold from refractory carbonaceous ores | |
| AU650355B2 (en) | Processing complex mineral ores | |
| WO2016181673A1 (en) | Mineral ore slurry pretreatment method, and method for manufacturing mineral ore slurry | |
| JP2010180450A (en) | Method for concentrating gold from copper-sulfide ore | |
| JP2019065341A (en) | Hydrometallurgical process for nickel oxide ore | |
| KR101352400B1 (en) | Method for producing copper concentrates from complex copper ore using leaching and precipitation | |
| JP2015214731A (en) | Gold recovery method | |
| Deng et al. | Treatment of oxidized copper ores with emphasis on refractory ores | |
| Dorfler et al. | Review of molybdenum recovery processes | |
| AU667635B2 (en) | Process for the recovery of silver by flotation from the residue from the wet extraction of zinc | |
| JP2012201920A (en) | Method for treating copper concentrate | |
| Lin | Characterization and flotation of sulfur from chalcopyrite concentrate leaching residue | |
| JP2011058018A (en) | Method for recovering gold concentrate from leach residue of copper sulfide minerals | |
| FI127007B (en) | Process for the treatment of sulphide ore |
Legal Events
| Date | Code | Title | Description |
|---|---|---|---|
| MK14 | Patent ceased section 143(a) (annual fees not paid) or expired |