PROCESS FOR THE ELECTROLYTIC RECOVERY OF COPPER FROM ITS SULFIDE ORES
United States Patent 3767543
An improvement in the ferric chloride leach recovery of copper from its sulfide ores which comprises recovery of the copper from the leach solution after removal of sulfur by electrolysis rather than by conventional cementation with added iron.
US Patent References:
METHOD FOR TREATING METALLIC SULFIDE COMPOUNDS
Brace - September 1969 - 3464904
Process of treating oxidized ores of lead
Christensen - August 1929 - 1726258
Process of treating lead-zinc sulphide ores
Christensen - May 1925 - 1539713
Process of treating sulphide ores of lead, silver, and copper
Christensen - March 1924 - 1485909
Process of treating ores containing galena
Christensen - May 1923 - 1456784
METHOD FOR TREATING METALLIC SULFIDE COMPOUNDS
Brace - September 1969 - 3464904
Process of treating oxidized ores of lead
Christensen - August 1929 - 1726258
Process of treating lead-zinc sulphide ores
Christensen - May 1925 - 1539713
Process of treating sulphide ores of lead, silver, and copper
Christensen - March 1924 - 1485909
Process of treating ores containing galena
Christensen - May 1923 - 1456784
Application Number:
05/157281
Publication Date:
10/23/1973
Filing Date:
06/28/1971
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Export Citation:
Assignee:
Hazen Research, Inc. (Golden, CO)
Primary Class:
Other Classes:
205/292, 205/112
International Classes:
C22B3/10; C25C1/00; C25C1/12; C22B3/00; C22D1/16; C23B5/20
Field of Search:
204/107,52R 75/117
US Patent References:
Other References:
Principles of Electroplating & Electroforming by Blum et al., 3rd ed.; 1949, pgs. 68-69. .
The Electromotive Series, "Simple Methods for Analyzing Plating Solutions," 7th ed, 1949, Hanson-Van Winkle-Munning Co., p. 20..
Primary Examiner:
Mack, John H.
Assistant Examiner:
Andrews R. L.
Claims:
I claim
1. The process for recovering copper from chalcopyrite concentrate which comprises:
2. The process of claim 1 in which the ferric chloride is recovered at the anode and is recycled to step (a).
3. The process of claim 1 in which iron in solution is recovered in a second electrolysis at the cathode and ferric chloride is recovered at the anode and recycled to step (a).
4. The process of claim 1 in which the pH of the leach solution subjected to electrolysis is maintained at about 2.3 or below.
5. The process of claim 1 in which the iron content of the leach solution subjected to electrolysis is up to about 180 grams per liter.
6. The process of claim 1 in which a cathode current density below about 20 asf is used for the electrolysis of copper to produce a coherent deposit on the cathode.
7. The process of claim 1 in which a cathode current density in excess of about 70 asf is used in the electrolysis of copper to provide a spongy deposit on the cathode.
1. The process for recovering copper from chalcopyrite concentrate which comprises:
2. The process of claim 1 in which the ferric chloride is recovered at the anode and is recycled to step (a).
3. The process of claim 1 in which iron in solution is recovered in a second electrolysis at the cathode and ferric chloride is recovered at the anode and recycled to step (a).
4. The process of claim 1 in which the pH of the leach solution subjected to electrolysis is maintained at about 2.3 or below.
5. The process of claim 1 in which the iron content of the leach solution subjected to electrolysis is up to about 180 grams per liter.
6. The process of claim 1 in which a cathode current density below about 20 asf is used for the electrolysis of copper to produce a coherent deposit on the cathode.
7. The process of claim 1 in which a cathode current density in excess of about 70 asf is used in the electrolysis of copper to provide a spongy deposit on the cathode.
Description:
SUMMARY OF THE INVENTION
In the present methods of treating copper sulfide ores for the recovery of copper the general practice involves smelting the sulfides and through a complex series of smelting operations driving off the sulfur as sulfur dioxide to produce a crude copper metal which is subsequently refined by electrolysis. This process is an ancient one which has been responsible for the production of most of the copper in the history of the world but suffers from some serious drawbacks. One very serious drawback of a smelting operation is that the sulfur dioxide produced from some steps of the operation is of such low grade that it is very uneconomical to produce sulfuric acid from it by means of the contact process and, accordingly, most smelting operations discharge this low grade sulfur dioxide waste gas into the atmosphere. This creates a serious pollution problem and for this reason the smelting process is the subject of intense investigative efforts at the present time to find ways of avoiding this pollution of the atmosphere.
In addition to the above disadvantage a smelting operation is economical only if done on a large scale and this accounts for the growth of central smelting facilities with the result that the capital investment for a smelting complex is large. Small mine operators who produce copper sulfide concentrate cannot afford the heavy capital investment of the pyrometallurgical smelters and must therefore sell the output of their mines to companies already possessing such facilities. This oftentimes leads to high shipping charges and high smelter toll fees.
Over the years many efforts have been made to find means of treating copper sulfide minerals by other methods than the ordinary smelting practice. Among these are processes in which the copper sulfide is first roasted to a mixture of copper oxide and copper sulfate with the production of by-product sulfur dioxide gas of a concentration high enough to permit its economic conversion into sulfuric acid.
Certain features of this process have disadvantages. The calcine from the furnace which contains the copper as a mixture of oxide and sulfate is leached with an acidic solution thereby producing a strong copper sulfate solution. This copper sulfate solution is then separated from the insoluble materials and subjected to electrolysis to produce copper metal and sulfuric acid. The latter is in such weak solution that it cannot be economically utilized, so that the loss of this by-product is a disadvantage. Furthermore, the disposal of the weak sulfuric acid solution presents an economic problem.
Another effort to solve the problem of recovery of copper values from copper sulfide minerals is the ferric chloride leaching method. This method has been studied at great length in various laboratories and particularly in the Canadian Department of Mines many years ago. (See "Investigations in Ore Dressing and Metallurgy 1924," Canada, Department of Mines, Mines Branch, John McLeish, Director, No. 643. ) In this process, the copper sulfide minerals are agitated with a hot solution containing a high concentration of ferric chloride. The ferric chloride acts as an oxidizing agent to attack the copper bearing sulfide minerals thereby converting the sulfide to elemental sulfur and putting the copper in solution as copper chloride. To the extent that iron is present in the sulfide mineral, as for example in chalcopyrite, this iron is dissolved as ferrous chloride and the associated sulfur is oxidized to elemental sulfur as is the case with the copper sulfide.
The ferric chloride leaching has very little effect upon the mineral pyrite when present by itself but, in general, will attack most sulfides such as pyrrhotite, chalcopyrite, chalcocite, sphalerite, and the like. Once the leaching reaction has taken place and the metallic elements are in solution as the chloride the sulfur remains in the residue as elemental sulfur. In this reaction the ferric chloride is reduced to ferrous chloride. The slurry is removed from the leaching vessel and sulfur is separated by any standard method such as filtration or countercurrent decantation thereby separating it from the metal. It can be seen that this is obviously a great advantage because the sulfur has been converted to a form that can be recovered and sold while the metallic elements remain in solution as chlorides for subsequent treatment.
In the process described in the Canadian Department of Mines publications, the leach solution containing copper chloride and ferrous chloride is treated by cementation in which metallic iron is added to the solution and the copper thereby precipitated as cement copper. This procedure, of course, increases the quantity of iron in solution which must be removed eventually and produces a copper of relatively low grade that must be purified and retreated. The solution remaining after electrolytic removal of copper and contained lead is then subjected to a second electrolysis where the iron is removed by deposition on a cathode in an electrolytic cell and some of the ferrous chloride in the anolyte is oxidized to ferric chloride, thus regenerating the leaching reagent for reuse.
In this flowsheet the amount of electrolytic iron which must be recovered is the total amount dissolved from the original concentrate feed and the amount which was added during the cementation for removal of the copper. The disadvantage of the recovery of the copper in this way is that the copper which is produced is of relatively low grade and is made expensive because of the cost of the iron which is added. In addition, this added quantity of iron must then be removed from the circuit by electrolysis in the next stage thereby requiring a considerable increase in the size of the electrolytic circuit that is required for the total iron removal with consequent increase in the cost for recovery of the copper.
In the flowsheet of this invention the teaching of the Canadian Department of Mines' process can be followed as far as the leaching and separation of liquids and solids but the improvement lies in the method of recovering the copper. It has been found, surprisingly, that it is possible to remove copper from the ferric chloride leach solution by electrolysis. Ordinary electrolysis of copper universally uses sulfuric acid solution as the electrolyte from which the copper is deposited on a cathode. It has been found that by conducting the electrolysis in a diaphragm-type cell, it is possible to remove the copper at quite high efficiency and high current density and at the same time generate ferric chloride in the anolyte compartment. This not only removes the requirement for the addition of iron in order to precipitate the copper but it regenerates the leaching reagent, ferric chloride, whereas the previous process penalized the circuit by the addition of iron above that in the concentrate which had to be removed subsequently. The net result of the present procedure is a significant reduction in the capital expense of a plant and the operating cost thereof for producing copper from sulfide minerals.
In ordinary electrowinning practice for copper a considerable effort is made to restrict the quantity of iron which is in solution with the copper. If the iron content of the solution becomes high (for example, 10g/1), a lowering of the current efficiency is observed and a change in the physical character of the deposited copper may take place. Also, if the iron content is very high, the current efficiency for the electrodeposition of the copper may drop to such a low value as to render the operation uneconomic. Therefore, the fact that the copper can be deposited from a solution which contains above 100 g/1 of iron as ferric chloride is quite surprising. The reason that the iron does not appreciably interfere with the current efficiency is probably because the electrolysis is done in a diaphragm cell so that the ferric iron which is formed at the anode cannot mix and come in contact with the deposited copper to redissolve it. Therefore, by performing the electrolysis in this manner, it is possible to take advantage of the electric current being used to deposit the copper and at the same time regenerate the leaching reagent without paying the penalty of low current efficiency ordinarily encountered when electrolysing a copper solution high in iron.
The iron in the solution after the electrolytic removal of copper is recovered by conventional electrolysis. This amount of iron is, of course, restricted to that which existed in the concentrate. Other metals are recovered by conventional means.
The process of the invention as illustrated in the accompanying drawing, a flow sheet of the process, will now be described, along with the equipment used.
The tests for which results are given below were performed using a ferrous chloride electrolyte medium having various adjusted concentrations of iron and to which cupric chloride was added in amounts to provide various concentrations of copper. The electrolyte medium simulated the leach solution which exists after copper sulfide concentrate is leached with ferric chloride. The equation representing the chemical reaction in the leaching step is:
4 Fe Cl 3 + Cu Fe S 2 ➝ Cu Cl 2 + 5 Fe Cl 2 + 2S.
After the sulfur is removed the electrolyte media contains essentially ferrous chloride and cupric chloride as does the electrolyte medium used in the tests.
A diaphragm cell was used in all tests. It consisted of a 1,500 cc cathode compartment separated from a 300 cc anode compartment by a Dynel filter cloth diaphragm. The anode compartment was equipped with a solution overflow. The anode was high purity graphite and had approximately a 0.1 sq.ft. area immersed. The cathode was a sheet of 22-gauge copper with approximately 0.33 sq.ft. submerged. During each of the tests, ferrous chloride electrolyte medium was continuously added to the cathode compartment. This solution flowed through the diaphragm into the anode compartment and out of the cell. The pH of the electrolyte was maintained at about 2.3 or below.
The electrolytic reactions are reduction of the copper to metal at the cathode and oxidation of ferrous iron to ferric at the anode. If insufficient ferrous iron is available for oxidation, chlorine is also produced at the anode.
In the first test the cell was filled with electrolyte medium which contained 65 g/1 copper and 135 g/1 iron. Approximately 100 cc/hr of 135 g/1 ferrous chloride solution was added to the cathode compartment and 12 amps were passed through the cell. After 8 hours of operation when practically all of the copper was plated out the test was terminated. The cathode had increased in weight by 116 grams. The voltage used was, of course, below the voltage requirement for the deposition of iron. The bulk of the cathode deposit was hard and coherent, but about 10 grams was spongy and contained only 70% copper. Data collected during the run are shown in the following table. ##SPC1##
The above results show that the process produces over 99 percent recovery of the copper in solution.
The iron in solution, of course, is electrolytically recovered in a subsequent stage by conventional electrolysis with simultaneous formation of ferric chloride which is recycled to the initial leach stage in a commercial process for treatment of copper sulfide concentrate.
Tests were run at various current densities to determine the most favorable current density ranges for forming spongy or coherent copper deposits, and to test the efficiency of the process as applied to low copper content electrolytes which correspond to electrolytes resulting from leaching low copper content concentrates.
In the second test the copper content of the used catholyte from the first test was increased to 2 g/1 copper by the addition of cupric chloride. The solution added to the cathode compartment during electrolysis contained 180 g/1 ferrous iron and 36 g/1 copper. The solution was added at approximately 160 cc/hr while 12 amps were passed through the cell. Approximately 36 grams of copper were added to the cell. At no time was the catholyte above 0.66 g/1 copper. Approximately all of the copper in solution was deposited at the cathode, half of which was coherent and the remainder spongy.
In the third test all conditions were the same as the second test except that 24 amps were used and the solution flow was increased to 335 cc/hr. In this test substantially all of the copper in solution was deposited as a spongy deposit.
The fourth test was also similar to those of the second test except that only 6 amps were used and the solution flow was reduced to only 85 cc/hr. Practically all of the copper deposited was in a coherent form.
The data from the tests where the catholyte copper composition was kept low could be summarized as follows: ##SPC2##
These results indicate that copper can be quantitatively recovered from concentrated ferrous chloride solutions (180 g/1 iron and above) by electrowinning. At current densities below about 20 asf the deposit is coherent and above 20 asf the deposit contains a gradually increasing amount of spongy material with practically all of it being spongy above about 70 asf. Copper solutions can be plated down to less than 0.1 g/1 copper. The percentage recovery of copper is not affected appreciatively by the concentration of the copper solution.
While the invention has been illustrated by its application to a flow sheet applied to calcopyrite, it is obviously not limited to this mineral as it is equally applicable to all concentrates of copper sulfide ores which can be leached with ferric chloride.
In the present methods of treating copper sulfide ores for the recovery of copper the general practice involves smelting the sulfides and through a complex series of smelting operations driving off the sulfur as sulfur dioxide to produce a crude copper metal which is subsequently refined by electrolysis. This process is an ancient one which has been responsible for the production of most of the copper in the history of the world but suffers from some serious drawbacks. One very serious drawback of a smelting operation is that the sulfur dioxide produced from some steps of the operation is of such low grade that it is very uneconomical to produce sulfuric acid from it by means of the contact process and, accordingly, most smelting operations discharge this low grade sulfur dioxide waste gas into the atmosphere. This creates a serious pollution problem and for this reason the smelting process is the subject of intense investigative efforts at the present time to find ways of avoiding this pollution of the atmosphere.
In addition to the above disadvantage a smelting operation is economical only if done on a large scale and this accounts for the growth of central smelting facilities with the result that the capital investment for a smelting complex is large. Small mine operators who produce copper sulfide concentrate cannot afford the heavy capital investment of the pyrometallurgical smelters and must therefore sell the output of their mines to companies already possessing such facilities. This oftentimes leads to high shipping charges and high smelter toll fees.
Over the years many efforts have been made to find means of treating copper sulfide minerals by other methods than the ordinary smelting practice. Among these are processes in which the copper sulfide is first roasted to a mixture of copper oxide and copper sulfate with the production of by-product sulfur dioxide gas of a concentration high enough to permit its economic conversion into sulfuric acid.
Certain features of this process have disadvantages. The calcine from the furnace which contains the copper as a mixture of oxide and sulfate is leached with an acidic solution thereby producing a strong copper sulfate solution. This copper sulfate solution is then separated from the insoluble materials and subjected to electrolysis to produce copper metal and sulfuric acid. The latter is in such weak solution that it cannot be economically utilized, so that the loss of this by-product is a disadvantage. Furthermore, the disposal of the weak sulfuric acid solution presents an economic problem.
Another effort to solve the problem of recovery of copper values from copper sulfide minerals is the ferric chloride leaching method. This method has been studied at great length in various laboratories and particularly in the Canadian Department of Mines many years ago. (See "Investigations in Ore Dressing and Metallurgy 1924," Canada, Department of Mines, Mines Branch, John McLeish, Director, No. 643. ) In this process, the copper sulfide minerals are agitated with a hot solution containing a high concentration of ferric chloride. The ferric chloride acts as an oxidizing agent to attack the copper bearing sulfide minerals thereby converting the sulfide to elemental sulfur and putting the copper in solution as copper chloride. To the extent that iron is present in the sulfide mineral, as for example in chalcopyrite, this iron is dissolved as ferrous chloride and the associated sulfur is oxidized to elemental sulfur as is the case with the copper sulfide.
The ferric chloride leaching has very little effect upon the mineral pyrite when present by itself but, in general, will attack most sulfides such as pyrrhotite, chalcopyrite, chalcocite, sphalerite, and the like. Once the leaching reaction has taken place and the metallic elements are in solution as the chloride the sulfur remains in the residue as elemental sulfur. In this reaction the ferric chloride is reduced to ferrous chloride. The slurry is removed from the leaching vessel and sulfur is separated by any standard method such as filtration or countercurrent decantation thereby separating it from the metal. It can be seen that this is obviously a great advantage because the sulfur has been converted to a form that can be recovered and sold while the metallic elements remain in solution as chlorides for subsequent treatment.
In the process described in the Canadian Department of Mines publications, the leach solution containing copper chloride and ferrous chloride is treated by cementation in which metallic iron is added to the solution and the copper thereby precipitated as cement copper. This procedure, of course, increases the quantity of iron in solution which must be removed eventually and produces a copper of relatively low grade that must be purified and retreated. The solution remaining after electrolytic removal of copper and contained lead is then subjected to a second electrolysis where the iron is removed by deposition on a cathode in an electrolytic cell and some of the ferrous chloride in the anolyte is oxidized to ferric chloride, thus regenerating the leaching reagent for reuse.
In this flowsheet the amount of electrolytic iron which must be recovered is the total amount dissolved from the original concentrate feed and the amount which was added during the cementation for removal of the copper. The disadvantage of the recovery of the copper in this way is that the copper which is produced is of relatively low grade and is made expensive because of the cost of the iron which is added. In addition, this added quantity of iron must then be removed from the circuit by electrolysis in the next stage thereby requiring a considerable increase in the size of the electrolytic circuit that is required for the total iron removal with consequent increase in the cost for recovery of the copper.
In the flowsheet of this invention the teaching of the Canadian Department of Mines' process can be followed as far as the leaching and separation of liquids and solids but the improvement lies in the method of recovering the copper. It has been found, surprisingly, that it is possible to remove copper from the ferric chloride leach solution by electrolysis. Ordinary electrolysis of copper universally uses sulfuric acid solution as the electrolyte from which the copper is deposited on a cathode. It has been found that by conducting the electrolysis in a diaphragm-type cell, it is possible to remove the copper at quite high efficiency and high current density and at the same time generate ferric chloride in the anolyte compartment. This not only removes the requirement for the addition of iron in order to precipitate the copper but it regenerates the leaching reagent, ferric chloride, whereas the previous process penalized the circuit by the addition of iron above that in the concentrate which had to be removed subsequently. The net result of the present procedure is a significant reduction in the capital expense of a plant and the operating cost thereof for producing copper from sulfide minerals.
In ordinary electrowinning practice for copper a considerable effort is made to restrict the quantity of iron which is in solution with the copper. If the iron content of the solution becomes high (for example, 10g/1), a lowering of the current efficiency is observed and a change in the physical character of the deposited copper may take place. Also, if the iron content is very high, the current efficiency for the electrodeposition of the copper may drop to such a low value as to render the operation uneconomic. Therefore, the fact that the copper can be deposited from a solution which contains above 100 g/1 of iron as ferric chloride is quite surprising. The reason that the iron does not appreciably interfere with the current efficiency is probably because the electrolysis is done in a diaphragm cell so that the ferric iron which is formed at the anode cannot mix and come in contact with the deposited copper to redissolve it. Therefore, by performing the electrolysis in this manner, it is possible to take advantage of the electric current being used to deposit the copper and at the same time regenerate the leaching reagent without paying the penalty of low current efficiency ordinarily encountered when electrolysing a copper solution high in iron.
The iron in the solution after the electrolytic removal of copper is recovered by conventional electrolysis. This amount of iron is, of course, restricted to that which existed in the concentrate. Other metals are recovered by conventional means.
The process of the invention as illustrated in the accompanying drawing, a flow sheet of the process, will now be described, along with the equipment used.
The tests for which results are given below were performed using a ferrous chloride electrolyte medium having various adjusted concentrations of iron and to which cupric chloride was added in amounts to provide various concentrations of copper. The electrolyte medium simulated the leach solution which exists after copper sulfide concentrate is leached with ferric chloride. The equation representing the chemical reaction in the leaching step is:
4 Fe Cl 3 + Cu Fe S 2 ➝ Cu Cl 2 + 5 Fe Cl 2 + 2S.
After the sulfur is removed the electrolyte media contains essentially ferrous chloride and cupric chloride as does the electrolyte medium used in the tests.
A diaphragm cell was used in all tests. It consisted of a 1,500 cc cathode compartment separated from a 300 cc anode compartment by a Dynel filter cloth diaphragm. The anode compartment was equipped with a solution overflow. The anode was high purity graphite and had approximately a 0.1 sq.ft. area immersed. The cathode was a sheet of 22-gauge copper with approximately 0.33 sq.ft. submerged. During each of the tests, ferrous chloride electrolyte medium was continuously added to the cathode compartment. This solution flowed through the diaphragm into the anode compartment and out of the cell. The pH of the electrolyte was maintained at about 2.3 or below.
The electrolytic reactions are reduction of the copper to metal at the cathode and oxidation of ferrous iron to ferric at the anode. If insufficient ferrous iron is available for oxidation, chlorine is also produced at the anode.
In the first test the cell was filled with electrolyte medium which contained 65 g/1 copper and 135 g/1 iron. Approximately 100 cc/hr of 135 g/1 ferrous chloride solution was added to the cathode compartment and 12 amps were passed through the cell. After 8 hours of operation when practically all of the copper was plated out the test was terminated. The cathode had increased in weight by 116 grams. The voltage used was, of course, below the voltage requirement for the deposition of iron. The bulk of the cathode deposit was hard and coherent, but about 10 grams was spongy and contained only 70% copper. Data collected during the run are shown in the following table. ##SPC1##
The above results show that the process produces over 99 percent recovery of the copper in solution.
The iron in solution, of course, is electrolytically recovered in a subsequent stage by conventional electrolysis with simultaneous formation of ferric chloride which is recycled to the initial leach stage in a commercial process for treatment of copper sulfide concentrate.
Tests were run at various current densities to determine the most favorable current density ranges for forming spongy or coherent copper deposits, and to test the efficiency of the process as applied to low copper content electrolytes which correspond to electrolytes resulting from leaching low copper content concentrates.
In the second test the copper content of the used catholyte from the first test was increased to 2 g/1 copper by the addition of cupric chloride. The solution added to the cathode compartment during electrolysis contained 180 g/1 ferrous iron and 36 g/1 copper. The solution was added at approximately 160 cc/hr while 12 amps were passed through the cell. Approximately 36 grams of copper were added to the cell. At no time was the catholyte above 0.66 g/1 copper. Approximately all of the copper in solution was deposited at the cathode, half of which was coherent and the remainder spongy.
In the third test all conditions were the same as the second test except that 24 amps were used and the solution flow was increased to 335 cc/hr. In this test substantially all of the copper in solution was deposited as a spongy deposit.
The fourth test was also similar to those of the second test except that only 6 amps were used and the solution flow was reduced to only 85 cc/hr. Practically all of the copper deposited was in a coherent form.
The data from the tests where the catholyte copper composition was kept low could be summarized as follows: ##SPC2##
These results indicate that copper can be quantitatively recovered from concentrated ferrous chloride solutions (180 g/1 iron and above) by electrowinning. At current densities below about 20 asf the deposit is coherent and above 20 asf the deposit contains a gradually increasing amount of spongy material with practically all of it being spongy above about 70 asf. Copper solutions can be plated down to less than 0.1 g/1 copper. The percentage recovery of copper is not affected appreciatively by the concentration of the copper solution.
While the invention has been illustrated by its application to a flow sheet applied to calcopyrite, it is obviously not limited to this mineral as it is equally applicable to all concentrates of copper sulfide ores which can be leached with ferric chloride.