Recovery of copper from chalcopyrite
United States Patent Pemsler [4 1 Apr. 29, 1975 RECOVERY OF COPPER FROM CHALCOPYRITE [75] Inventor: J. Paul Pemsler, Lexington. Mass.  
 [73] Assignee: Kennecott Copper Corporation,  
 New York, NY.  
 [22] Filed: Jan. 28, 1974 [2]] Appl. No.: 437,237  
 [52] US. Cl. 75/72; 75/21; 75/117 [51] Int. Cl C22b 15/00 [58] Field of Search 75/21, 72, 117; 204/106 [56] References Cited UNITED STATES PATENTS 3,681,055 8/1972 Little ct a1. 204/106 3,761,245 9/1973 Bingham 75/21 3,799,764 3/1974 Opic et ul. 75/72 SALT RECYCLE Primary E.\&#39;aminerL. Dewayne Rutledge Assistant Examiner-M. J. Andrews Attorney, Agent, or Firm-John L. Sniado; Lowell H. McCarter; Anthony M. Lorusso [57] ABSTRACT Copper concentrate is mixed with iron and a fusible salt or fusiblesalt mixture which is molten at the roasting temperature. The mixture is roasted in a nonoxidizing atmosphere at a temperature below 475C (preferably between 350C390C) to produce a product containing free copper and pyrrhotite. The salt acts as a solvent for the reaction and enables copper to be recovered economically from copper concentrate without the evolution of sulfide pollutants into the atmosphere.  
 44 Claims, 5 Drawing Figures F6 RECYCLE CONCENTRATE STORAGE fizz SALT MAKE-UP SA LT FURNACE SLURRY QUENCH FLOTATION THICKENER HS TO TAILINGS DUMP c c.o WASH MELTING FURNACE FIJENTEBAPRZSBYS 3,880,650  
 sum 1 BF 3 SALT RECYCLE Fe RECYCI I Fe 4 2s /22 SALT MAKE-UP l8 SALT I6 20 I0 l2 32 CONCENTRATE STORAGE 3o WATER H m \-J Y 46 l 2 44 FURNACE 52 C SLURRYQUENCH SALT POND 0.6.0 I, l Cu WASH FLOTATION 70 THICKENER FeS TO TAILINGS DUMP  72 ANODE MELTING FURNACE I00- F/G.2.  
 73 CONVERSION 60 l l l l I l I l V I TEMPERATURE, *0 FIG. 3.  
 IOO-I CONVE RSION l I I l I I I l I 0 TIME, HRS.  
 PQJENTEBAPRZSIQYS 880.650  
 SHEET 3 BF 3 F76 so WAEIGHT PERCENT s Cu ISOTHERMAL SECTION AT 400C Fe WEIGHT PER CENT $(LIQUID) Cu ISOTHERMAL SECTION AT 500C Fe WEIGHT PER CENT RECOVERY OF COPPER FROM CHALCOPYRITE BACKGROUND OF THE INVENTION One of the worlds major sources of copper is the ore chalcopyrite (CuFeS The traditional method for recovering copper from copper concentrates which are comprised principally of chalcopyrite is by smelting. In the conventional smelting of copper sulfide concentrates. a copper-iron-sulfide matte is produced in a reverberatory-type furnace. The molten matte is then transferred to converters, where, in the first step of a batch operation, the iron sulfide is oxidized to yield sulfur dioxide and an iron oxide. The iron oxide is reacted with silica flux to form a slag. In a second step, the copper sulfide is oxidized to yield copper and sulfur dioxide. Thus, the smelting method produces sulfur containing gases which present an onerous air pollution problem.  
  In view of the present day concern over the quality of the air, a significant effort has been made in an attempt to produce copper without polluting the air. However, an air pollution free process for recovering copper from copper concentrate which is also economically competitive with present day smelting methods has not materialized. One problem which has plagued the industry is the high capital cost associated with equipment for removing sulfur containing gases from the smelter before they reach the atmosphere.  
 SUMMARY OF THE INVENTION The process of the present invention produces copper from copper concentrates without the evolution of sulfur containing gases. Furthermore, the equipment needed to recover copper in accordance with the present invention is relatively inexpensive when compared with conventional smelting equipment. In accordance with the present invention, iron and a fusible salt or fusible salt mixture is mixed with the ore and the mixture is roasted at a temperature below 475C. As a result the copper separates from the concentrate without the evolution of sulfur containing gases.  
  Accordingly, it is an object of the invention to provide a process for the recovery of copper from copper concentrate which does not present a serious air pollution problem.  
  A further object to the present invention is to provide a process for the recovery of copper from copper concentrates which utilized equipment which is comparatively inexpensive.  
  A further object of the present invention is to provide a process in which chalcopyrite is roasted to remove the sulfur without the evolution of sulfur containing gases.  
  Yet another object of the present invention is to provide a process for the recovery of copper and other metals from copper concentrates which includes the step of roasting the copper concentrate with iron and a fusible salt medium at a temperature below 475C to decompose the sulfide ore in the copper concentrate and produce a reaction product containing an iron sulfide and a metal phase containing copper and other valuable metals.  
 BRIEF DESCRIPTION OF THE DRAWING FIG. 1 is a flow chart illustrating the process of the present invention,  
  FIG. 2 is a graph illustrating the effect of temperature on the process of the present invention,  
  FIG. 3 is a graph illustrating the percent conversion of concentrate versus time,  
  FIG. 4 is a phase diagram of copper, sulfur and iron at 400C, and  
  FIG. 5 is a phase diagram of copper, sulfur and iron at 500C.  
 DESCRIPTION OF THE PREFERRED EMBODIMENTS In accordance with the present invention, copper concentrate is mixed with a fusible ionic salt mixture and iron and is roasted in a non-oxidizing atmosphere at a temperature below 475C. Preferably the mixture is roasted at a temperature between 350C and 390C. The product ofthe reaction is free copper and pyrrhotite (FeS). To insure a high yield of copper an excess amount of iron over the stoichiometric amount required is employed. The roasted product is quenched and fed to a magnetic separator to remove excess magnetic iron. After suitable solid-liquid separation, a floatation is performed to provide a copper product. The copper product is fed to a melting furnace for further refining.  
  The major constituent of copper concentrate is chalcopyrite which has the formula CuFeS As is stated above, a significant problem associated with the smelting of chalcopyrite is that the sulfur therein forms a sulfide gas which pollutes the atmosphere. In accordance with the present invention, no significant amount of sulfur containing gases are produced when the copper in the chalcopyrite is separated from the remainder of the compound. This separation proceeds in accordance with the following reaction:  
  The molten salt acts as a solvent which enables a separation of the chalcopyrite in accordance with the following mechanism:  
  As is shown aboye, the mechanism of the reaction involves an ionic diffusion process wherein chalcopyrite partially dissolves in the molten salt. Sulfide ions thus produced react with iron releasing electrons which may, in turn, reduce copper ions to copper metal. The use of a molten salt solvent or catalyst enables a rapid reaction between metallic iron and chalcopyrite to directly produce metallic copper. Furthermore, because the sulfur in the chalcopyrite immediately reacts with the iron, no sulfur containing gases are emitted. Further details relative to the molten salt medium appear below.  
  The process of the present invention can be advantageously employed on all known copper ores and concentrates containing copper and sulfur.  
  The term copper concentrate&#34; is meant to include copper sulfide concentrates, whether produced from ore by flotation or produced as a sulfide matte by smelting. The chief ingredients of a copper-iron sulfide concentrate or matte are copper, iron and sulfur. Copper may range from 15 to 45 percent, preferably, 18 to 32 percent; iron from 15 to 40 percent, more preferably, 20 to 35 percent; and sulfur from 20 to 45 percent, more preferably, 25 to 40 percent; and the balance some gangue material, e.g. SiO A1 etc. The foregoing concentrates may contain small amounts of other ingredients, for example, one or more of Se, Te, Pb, Zn and such previous metals as Au, Ag, Pt and Pd. Low grade sulfide ores consisting mostly of chalcopyrite (commonly written as CuFeS more rarely as Cu S, Fe S disseminated as minute specks constituting a very small percentage of the weight of the host rock. Those portions of the mineralized deposits commonly regarded as commercial ores usually assay 0.35 to 0.85 percent copper. Theoretically, chalcopyrite consists of 34.5 percent copper, 30.5 percent iron and 35.0 percent sulfur but actually there is usually a small content of iron pyrite, FeS blebs so finely intermixed that perfect separation is impossible. Occurrences of other copper sulfide minerals are known such as chalococite, Cu S, bornite, Cu FeS covellite, CuS, and enargite, Cu AsS,, but chalcopyrite, CuFeS is by far the most common mineral of copper as found in nature. The ore, as mined, is crushed in several stages, ground to suitable size and concentrated by flotation processes to a product consisting, typically, of 28 percent copper, 30 percent iron, 32 percent sulfur, 9 percent insoluble (SiO etc.) and 1 percent miscellaneous (CaCO etc.). These typical analyses represent determination made on a dry basis. Actually, the concentrator product is usually a filter cake containing 9-10 percent residual water which may or may not be dried before shipment to a smelter.  
  Other sulfides behave similarly to chalcopyrite when roasted in accordance with the teachings of the present invention. For example separation of copper from other sulfides proceeds as follows:  
  In order to separate copper from the sulfide in accordance with the reaction set forth above, it is necessary that the roasting with iron be performed at temperatures below 475C. Indeed, the literature in this art, suggests that the foregoing reactions should occur at temperatures as high as 475C. However, little yield results at temperatures above 420C. Furthermore, in accordance with the present invention the most favorable yield results at temperatures between 350C and 390C.  
  Two isothermal sections of a copper-iron-sulfide system are set forth in FIGS. 4 and 5. FIGS. 4 and 5 are taken from an article entitled Thermal Stability of Assemblages in the Cu-Fe-S System by R. A. Yund and G. Kullerud which appears in the Journal of Petrology, Vol. 7, Part 3, Page 454-488 (1966), the teachings of which are incorporated herein by reference. In FIGS. 4 and 5,  
 W Py  
 po pyrrhotite cp chalcopyrite ch cubanite id idaite bn bornite cov covelite cch chalcocite.  
  FIG. 5 indicates that if iron is added to chalcopyrite at 500C, the chalcopyrite would be converted to bornite and pyrrhotite. Further addition of iron would not result in any additional desulfurization or conversion to metallic copper. FIG. 4, however, indicates that a major tie line change occurs between 400C and 500C. FIG. 4 also shows that the addition of iron to chalcopyrite or bornite will yield pyrrhotite and copper.  
  The following is a description of the overall aspects of the process of the present invention.  
  As is shown in FIG. 1, concentrate 10 is conveyed by a conveyer 12 to a storage bin 14. As the concentrate is required, it is fed from storage into a hopper 16, as is shown by arrow 18. Iron is loaded into a hopper 20, as is shown by arrow 22. The fusible salt mixture is fed into a hopper 24 as is shown by arrow 26.  
  Hoppers 16, 20, 24 regulate the flow of concentrate, iron and salt. These constituents can be mixed on a conveyer 28. The mixed constituents are delivered to another conveyer 30 which loads the mixed constituents into a storage hopper 32.  
  A gondola 34 is positioned beneath storage hopper 32. The go ndola is filled with constituents from hopper 32 as is shown by arrow 36. After gondola 36 is filled, it enters a furnace 38. As is shown in FIG. 1, gondola 34 is part of a train. When one gondola in the train is filled, it is advanced and the next gondola in the train is positioned under hopper 32.  
  The temperature of the furnace 38, is about 375C and a loaded gondola is in the furnace for about one hour. After leaving the furnace, the reaction product in the gondolas are dumped into a well 40. The reaction product is quenched by water which is flowed into well 40 and is agitated by a mixer to produce a slurry which is withdrawn as is shown by arrow 44. The water also dissolves the salt. The slurry contains particles of copper, iron, iron sulfide and gangue.  
  The slurry from well 40 is delivered to a magnetic separator 46.. The magnetic separator 46 removes metallic iron present in excess of the stoichiometric amount from the slurry. This iron is then recycled along route 48 and is also fed into hopper 20.  
  The slurry leaving the magnetic separator 46 is fed into a countercurrent decantation wash system 50 (CCD wash) as is shown by arrow 52. In vessel 54 the solids are separated from the remainder of the salt water solution and are passed into vessel 56. The salt water solution is withdrawn from vessel 54 and is delivered to a salt pond as is shown by arrow 58. In the salt pond the water is evaporated to produce salt for a recycle. Recovered salt is recycled along route to hopper 24.  
  After being washed in vessel 56, the metal values are delivered into another wash vessel 62. After being washed in vessel 62, the metal values are delivered to a floatation cell 64. Fresh water is introduced into vessel 62, is flowed counter-current to the solids into vessel 56 as is shown by arrow 66. After flowing through vessel 56, the water is flowed into well 40 to be used to quench the reaction product as is shown by arrow 68.  
  The metal values recovered from vessel 62 as is shown by arrow 70, can be recovered in any conventional manner. For example, the metal values may be recovered in a floatation cell. The metal values (principally copper) are sent to a melting furnace and then to electro refining as is shown by arrow 72. The tailings which consist mainly of FeS go to a waste dump.  
  Again it should be noted that the process of the present invention involves the separation of sulfur from copper sulfide ores without the evolution of sulfur containing gases. To accomplish the foregoing, the sulfur is separated from the copper in a molten salt medium. The molten salt acts as a solvent for the constituents in the chemical reaction or in a broader sense as a catalyst for the reaction. Many salts and mixtures of salts can be advantageously employed in the process of the present invention as the solvent for the reaction. The requirements of the solvent are that it possesses the following properties:  
 1. non-oxidizing;  
 2. molten at at the roasting temperature;  
 3. exhibit ionic characteristic to act as a solvent for ore:  
 4. capability to dissolve sulfide ions: and  
 5. unreactive with the reactants and products of the reaction.  
  Various fused salt mixtures have been used in the study of this invention and it has been found that cer tain eutectic mixtures of ionic salts possess all of the foregoing characteristics. The simplest mixture that possesses these characteristics is the lithium chloridepotassium chloride eutectic. Since the lithium salt is expensive; however, a commercialized process would not make use of this material. Chloride eutectics are preferred because they are inexpensive to prepare. The preferred chloride eutectic is a combination of sodium chloride, potassium chloride and magnesium chloride. Other mixtures which are satisfactory are aluminum chloride-sodium chloride; iron chloride-sodium chloride, and those non-oxidizing mixtures with melting points below 475C (preferable below 390C), appearing in Chapter I of Molten Salts Handbook, by George J. .lanz, Academic Press, New York, 1967, the teachings of which are incorporated herein by reference.  
  The fine particulate iron is readily available from a number of commercial sources. Ungraded 200 mesh iron powder of the type commonly used for coating welding rods can be advantageously employed in the process of the present invention.  
  As is set forth above, the roasting is performed in a non-oxidizing atmosphere. As used throughout the specification and claims, the term non-oxidizing atmosphere&#34; means an atmosphere in which metallic iron will not be oxidized. At the temperature range employed in this process. this is equivalent to maintaining the oxygen partial pressure below a value of l0&#39; atm. This may be achieved by utilizing any of a number of so-called reducing gases commonly used in industry in annealing and other similar operations. Typically, these are produced by incomplete combustion of any fuel materials such as oil, coal, natural gas or propane. Those skilled in the art are well familiar with such techniques.  
  Various factors affecting the yield have been investigated. FIG. 2 shows that the effect of temperature on the percent conversion under identical conditions of feed material and time. To prepare the curve of FIG. 2, a 200 mesh concentrate having the composition set forth in Table I below was reacted with -200 mesh iron and a 44 weight percent LiCl 56 weight percent KCl eutectic. The reaction time and feed rate were fixed.  
 Reaction rates peak over about a 40 temperature range from 350C to 390C. The rate is believed to fall off at higher temperatures due to the approach of the tie line change wherein copper metal would no longer be in thermodynamic equilibrium with pyrrhotite. The falloff on the low side is believed to be due to the normal decreases of rate with temperature and/or changes in the solubility of the minerals in the fused salt with temperature.  
  FIG. 3 shows a plot of the percent conversion of concentrate versus time at 375C using 200 mesh iron, a 44 weight percent LiCl- 56 weight percent KCI eutectic, and a 200 mesh concentrate having the composition set forth in Table I below.  
 TABLE I Analysis of typical concentrate in percent by weight Copper 26 9 Molybde- 0.l67 Alumi- 0.200  
 num num Iron 27.1 Silver 0.020 Silicon 2.46  
 Sulfur 30.0 Calcium 0.230  
  It can be seen that over percent conversion of concentrate to metallic copper occurs within one hour. The rate has also been shown to vary with particle size of both concentrate and iron. Deliberately screened coarser concentrate fractions react somewhat slower. The use of finer iron powder increases the rate. Conversions in excess of percent can be obtained in time substantially shorter than one hour.  
  The reaction product is a fine dispersion of whiskerlike copper in a mixture of FeS, gangue minerals, and excess iron. Mass spectographic analysis of impurities in copper produced by conversion of concentrate indicates a purity level of 99.90 percent copper exclusive of silver and gold. The results in parts per million by weight (ppm W) are presented in the Table II below.  
 TABLE II IMPURITY ANALYSIS Copper may be readily leached from the reaction product, using ammonia-CO solutions. With a 0.6M CO M NI-I aqueous solution, 95 percent of the converted copper is leached out within 15 minutes. Thus, in accordance with the present invention, copper can be recovered from the reaction product by leaching and electrowinning. This is advantageously accomplished by leaching with aerated ammonia-ammonium carbonate solutions and subsequently electrowinning.  
  It is recognized, however, that leaching and electrowinning are not necessarily the most economical route for the recovery of the copper. Since the copper is present as metallic copper, it is desirable to physically separate the copper from the mixture and directly melt it to cathodes for refining. Thus, it is preferred to separate the copper by floatation. Enrichment from 13.5 percent by weight Cu in the reaction product to 35 40 percent by weight Cu was achieved in a single floatation.  
  Atomic absorption analysis of the copper metal and original concentrate indicate that all of the silver and gold report to the copper, while all the molybdenum remains with the pyrrhotite. The product could therefore be electrorefined to reclaim precious metals.  
  Magnetic separation studies with a Davis tube indicated that substantially all of the excess reactant iron is readily recovered by the procedure set forth above. Some entrapped copper and pyrrhotite were present in the iron but this presents no problem since they are recycled with the iron.  
  The process of the present invention is further illustrated by the following non-limiting example.  
 EXAMPLE l Referring to FIG. 1, 647 tons per day of 200 mesh copper concentrate is loaded into hopper 16. The concentrate has the average composition as set forth in Table 1 above.  
  Along with the copper concentrate, l92 tons per day of iron and 40 tons per day of salt is introduced into hopper 20 and 24 respec ively. In connection with the foregoing amounts of iron and salt it should be noted that the foregoing tonage does not include the amount of iron and salt that is recycled. The iron added into hopper 20 is 200 mesh iron. The salt is a eutectic mixture of 55 weight percent MgCl 25 weight percent NaCl and 20 weight percent KCl. After being mixed, the concentrate, iron, and salt is batch heated in a furnace at a temperature of 390C for 1 hour. To maintain a non-oxidizing atmosphere, a commercial fuel such as coal or oil is incompletely combusted.  
  Immediately after being roasted in the furnace, the reaction product is quenched with water to form a slurry. The slurry is then passed through a magnetic separator to remove excess iron. 42 tons per day of iron is recycled back to hopper 20 from the magnetic separator. After being subjected to magnetic separation, the slurry is subjected to a countercurrent decantation wash. in which the salt water solution is withdrawn from the remaining solid particles. The water is evaporated to leave dry salt. 360 tons per day of recycled salt is fed into hopper 24 along with the salt make up. Thus, the total amount of salt required to treat 647 tons of concentrate is 400 tons while the total amount of iron required is 234 tons. The water required in the countercurrent decantation wash is about 43,000 gallons per day. After floatation, about l5l.8 tons of copper per day is withdrawn from the melting furnace.  
  In view of the foregoing teachings of the present invention, it is possible to remove sulfur from a sulfide copper ore without causing an appreciable air pollutant in the form of a gas containing a sulfide. This is made possible by using a molten salt medium as a catalyst or solvent for the reaction. Variations in the parameters disclosed, however, are well within the skill of those in this art in view of the teachings of the present invention.  
  Thus, the invention may be embodied in other specific forms without departing from the spirit or essential characteristics thereof. The present embodiments are therefore to be considered in all respects as illustrat ive and not restrictive, the scope of the invention being indicated by the appended claims rather than by the foregoing description, and all changes which come within the meaning and range of equivalency of the claims are therefore intended to be embraced therein.  
 I claim:  
  1. A process for recovering copper from a concentrate containing copper and sulfur comprising the following steps:  
 a. mixing the concentrate with iron and a fusible ionic salt;  
 b. roasting the mixture of step (a) at a temperature below 475C to produce a reaction product containing elemental copper and pyrrhotite; and  
 c. separating the copper from the reaction product.  
  2. The process as set forth in claim 1 wherein the roasting which takes place in step (b) is performed in a non-oxidizing atmosphere.  
  3. The process as set forth in claim 2 wherein in step (a), an excess amount of iron over the stoichiometric amount required is mixed with the concentrate and salt.  
  4. The process as set forth in claim 1 wherein prior to separating the copper from the reaction product, the reaction product is quenched in water to dissolve the ionic salt and produce a slurry.  
  5. The process as set forth in claim 4 also including the step of feeding the reaction product to a magnetic separator to remove magnetic iron from the reaction product.  
  6. The process as set forth in claim 5 wherein in step (c) the copper is separated from the iron by floatation.  
  7. The process as set forth in claim 6 wherein prior to floatation the water quenched reaction product is subjected to a solid-liquid separation.  
  8. The process as set forth in claim 7 wherein the copper that is separated is refined in a melting furnace.  
  9. The process as set forth in claim 1 wherein in step (b) the mixture is roasted at a temperature of between about 350C and 390C for one hour.  
  10. The process as set forth in claim 5 wherein the magnetic iron removed in the magnetic separator is recycled and is mixed with the concentrate and fusible salt in step (a).  
  11. The process as set forth in claim 10 wherein water and dissolved ionic salt are withdrawn from the slurry to leave solids and the ionic salt is recovered and recycled to be mixed with the concentrate and iron in step (a).  
  12. The process as set forth in claim 11 wherein the solids are washed with water in a countercurrent decantation wash.  
  13. The process as set forth in claim 12 wherein the water used to wash the solids is cycled after flowing through the countercurrent decantation wash to quench the reaction product.  
  14. The process as set forth in claim 1 wherein in step (a) the ionic salt that is mixed with the iron and concentrate has the following properties:  
 1. is non-oxidizing;  
 2. is molten at temperatures of 475C or lower;  
 3. exhibits ionic characteristics to act as a solvent for the concentrate;  
 4. is capable of dissolving sulfide ions; and  
 5. is unreactive with the reactants and products of the reaction.  
  15. The process as set forth in claim 14 wherein a eutectic mixture of two or more salts is the ionic salt that is mixed with the iron and concentrate in step (a).  
  16. The process as set forth in claim 1 wherein a chloride eutectic is the ionic salt that is mixed with the iron and concentrate in step (a).  
  17. The process as set forth in claim 1 wherein a mixture of sodium chloride, potassium chloride and magnesium chloride is the ionic salt that is mixed with the iron and concentrate in step (a).  
  18. The process as set forth in claim 1 wherein a eutectic mixture of sodium chloride, potassium chloride and magnesium chloride is the ionic salt that is mixed with the iron and concentrate in step (a).  
  19. The process as set forth in claim 1 wherein a member of eutectic salts selected from the group consisting of aluminum chloride-sodium chloride and iron chloride-sodium chloride is the ionic salt that is mixed with the iron and concentrate in step (a).  
  20. The process as set forth in claim 1 wherein in step (b) the mixture is roasted at a temperature between the range of 350C390C.  
  21. The process as is set forth in claim wherein the iron that is mixed with the concentrate and ionic salt in step (a) is particulated to a size of about -200 mesh.  
  22. The process as set forth in claim 1 wherein in step (c) the copper is separated from the reaction product by leaching the reaction product with an ammonia ammonium carbonate solution and subsequently electrowinning the copper.  
  23. A process for recovering copper from a material containing copper and sulfur, the material including a member selected from the group consisting of chalcopyrite, bornite, chalcocite, covellite, pyrite, cubanite and idaite comprising the following steps:  
 a. mixing the material with iron and a fusible ionic salt; the salt being, non-oxidizing, molten at temperatures of 475C or lower, a solvent for the material, capable of dissolving sulfide ions, and unreactive with the reactants and products of the reaction;  
 b. roasting the mixture of step (a) at a temperature below 475C to partially dissolve the material into the salt and to produce a reaction product containing elemental copper and pyrrhotite, said salt acting as a solvent enabling the reduction of copper by iron to occur below 475C without the evolution of sulfide pollutants into the atmosphere, and  
 c. separating the copper from the reaction product.  
  24. The process as set forth in claim 23 wherein the material includes chalcopyrite.  
  25. The process as set forth in claim 24 wherein the roasting which takes place in step (b) is performed in a non-oxidizing atmosphere.  
  26. The process as set forth in claim 25 wherein in step (a), an excess amount of iron over the stoichiometric amount required is mixed with the material and salt.  
  27. The process as set forth in claim 24 wherein prior to separating the copper from the reaction product, the reaction product is quenched in water to dissolve the ionic salt and produce a slurry.  
  28. The process as set forth in claim 27 also including the step of feeding the reaction product ,to a magnetic separator to remove magnetic iron from the reaction product.  
  29. The process as set forth in claim 28 wherein in step (c) the copper is separated from the iron by floatation.  
  30. The process as set forth in claim 29 wherein prior to floatation, the water quenched reaction product is subjected to a solid-liquid separation.  
  31. The process as set forth in claim 30 wherein the copper that is separated is refined in a melting furnace.  
  32. The process as set forth in claim 24 wherein in step (b) the mixture is roasted at a temperature of between about 350C and 390C for one hour.  
  33. The process as set forth in claim 28 wherein the magnetic iron removed in the magnetic separator is recycled and is mixed with the material and fusible salt in step (a).  
  34. The process as set forth in claim 33 wherein water and dissolved ionic salt are withdrawn from the slurry to leave solids and the ionic salt is recovered and recycled to be mixed with the material and iron in step (a).  
  35. The process as set forth in claim 34 wherein the solids are washed with water in a countercurrent decantation wash.  
  36. The process as set forth in claim 35 wherein the water used to wash the solids is cycled after flowing through the countercurrent decantation wash to quench the reaction product.  
  37. The process as set forth in claim 24 wherein a eutectic mixture of two or more salts is the ionic salt that is mixed with the iron and material in step (a).  
  38. The process as set forth in claim 27 wherein a chloride eutectic is the ionic salt that is mixed with the iron and concentrate in step (a).  
  39. The process as set forth in claim 23 wherein a mixture of sodium chloride, potassium chloride and magnesium chloride is the ionic salt that is mixed with the iron and material in step (a).  
  40. The process as set forth in claim 24 wherein a eutectic mixture of sodium chloride, potassium chloride and magnesium chloride is the ionic salt that is mixed with the iron and material in step (a).  
  41. The process as set forth in claim 24 wherein a member of eutectic salt selected from the group consisting of aluminum chloridesodium chloride and iron chloridesodium chloride is the ionic salt that is mixed with the iron and material in step (a).  
  42. The process as set forth in claim 23 wherein in step (b) the mixture is roasted at a temperature between the range of 350C390C.  
  43. The process as set forth in claim 43 wherein the iron that is mixed with the material and ionic salt in step (a) is particulated to a size of about 200 mesh.  
  44. The process as set forth in claim 23 wherein in step (c) the copper is separated from the reaction product by leaching the reaction product with an ammonia ammonium carbonate solution and subsequently electrowinning the copper.