CA1258181A - Process for producing enriched mineral ore concentrates - Google Patents
Process for producing enriched mineral ore concentratesInfo
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- CA1258181A CA1258181A CA000477505A CA477505A CA1258181A CA 1258181 A CA1258181 A CA 1258181A CA 000477505 A CA000477505 A CA 000477505A CA 477505 A CA477505 A CA 477505A CA 1258181 A CA1258181 A CA 1258181A
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- desired metal
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- copper
- mineral ore
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- 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
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Abstract
ABSTRACT
A process for selectively enriching a sulfide mineral ore concentrate containing desired metal sulfides and mixed metal sulfides containing the desired metal and other metal or other metal sulfides, the other metal having a lower affinity for sulfur than the desired metal to produce an enriched concentrate having an increased desired metal content;
the process comprising:
(a) reacting a desired metal ion rich aqueous stream with the sulfide mineral ore concentrate in a reaction zone to produce a product aqueous slurry containing increased quantities of desired metal sulfides;
(b) separating the product aqueous slurry into a desired metal ion depleted aqueous stream and a finely divided particulate solids stream containing the desired metal sulfides; and, (c) separating at least a portion of the desired metal sulfides from the finely divided particulate solids stream by a froth flotation process to produce the enriched concentrate.
A process for selectively enriching a sulfide mineral ore concentrate containing desired metal sulfides and mixed metal sulfides containing the desired metal and other metal or other metal sulfides, the other metal having a lower affinity for sulfur than the desired metal to produce an enriched concentrate having an increased desired metal content;
the process comprising:
(a) reacting a desired metal ion rich aqueous stream with the sulfide mineral ore concentrate in a reaction zone to produce a product aqueous slurry containing increased quantities of desired metal sulfides;
(b) separating the product aqueous slurry into a desired metal ion depleted aqueous stream and a finely divided particulate solids stream containing the desired metal sulfides; and, (c) separating at least a portion of the desired metal sulfides from the finely divided particulate solids stream by a froth flotation process to produce the enriched concentrate.
Description
A PROCESS FOR PRODUCING ENRICHED
MINERAL ORE CONCENTXATES
Field of the Invention This invention relates to the selective enrichment of mineral ore concentratesn This invention further relates to the enrichment of mineral ore concentrates by a combination of a replacement step to produce additional quantities of sulfides of the desired metal in the concentrate solids and a selective froth flotation step to selectively recover the desired metal sulfides.
Description of the Prior Art In many commercial ore deposits, particularly those of copper, molybdenum, nickel, zinc and silver, the ore consists of a mixture of associated metal sulfide minerals.
Numerous minerals can be associated in any one orebody, and the specific minerals associated will vary from deposit to deposit. A large numbez of typical associations are described in Peters, Exploration and Mininq Geology (1978) Table 2-1.
For instance, in copper deposits in Utah and Chile, the copper-containlng minerals Chalcopyrite (CuFeS2), ~,r~
--1-- ~
bornite (cusFes4) and sometimes enargite (Cu3AsS4) are associated with pyrite (FeS2), molybdenite (MoS2) and in some cases sphalerite (ZnS) and galena (PbS). Similarly, a Montana copper deposit contains the copper-bearing minerals chalcopyrite, chalcocite (Cu2S), bornite, enargite, covellite (CuS) and sometimes tetrahedrite (CU3SbS3) and tennantite (Cu3Ass3) in association with pyrite and in some instances molybdenite, sphalerite and galena Such associations are not limited only to the copper deposits. Molybdenum sulfides are found in association with copper sulfides as noted above, and also with pyrite and magnetite (Fe3o4). Similarly, silver sulfides such as argentite (Ag2S) and polybasite (AgSbS6) are found in association with galena, chalcopyrite, pyrite and some tellurides. Nickel sulfides, such as pentlandite (Fe,Ni)gsg, are found in association with pyrrhotite (FeS), pyrite, chalcopyrite, marcasite (FeS2), magnetite and nickel arsenides. Zinc sulfides, such as sphalerite, are found associated with pyrrhotite, galena, chalcopyrite, pyrite and magnetite. In addition, it is common for some copper ores to contain small amounts of precious metal minera~s and trace metals such as bismuth, selenium and tellurium. It will be understood by those skilled in the art that the minerals listed above and mentioned hereinafter are described in idealized chemical form for convenience. It will be recognized, however, that most field ores actually have somewhat variable mineral composition and may include such minerals to a variable degree. See, for instance, E. S. Dana, A Textbook of Mineralogy, p 354 (4th ed., 1932).
1258~1 It will be noted from the preceding list that the deposits contain mixtures of both binary minerals such as chalcocite and covellite and ternary minerals such as chalcopyrite and bornite. More comple~ minerals may also be present.
In conventional mineral processing practice, it has been common to concentrate the ore containing metal values by processes such as grinding and flotation. The concentrate is then fed to a metal extraction process such as smelting to recover the desired metal values. The principal purpose of conventional ore concentration is two-fold: ~a) to separate minerals containing the desired metal from gangues (waste materials) to avoid having the gangues go through the desired metal separation process; and, (b) to separate those minerals containing the desired metal from minerals containing other desirable metals. Concentration is also beneficial when the concentrate of desired metal values is to be transported a substantial distance for smelting or other processing.
Increasing the weight percent desired metal in the concentrate reduces the weight and shipping costs of the concentrate per unit of desired metal in the concentrate.
The amount of upgrading possible by conventional concentration processes is limited by the chemical composition of the minerals containing the metal to be recovered. Further upgrading requires a change in the chemical composition of the minerals. There are two methods for enriching chalcopyrite copper concentrates by eliminating iron and sulfur. The first is conventional reverberatory roasting which eliminates iron in slag and sulfur as SO2. The second method involves 1~Z581~1 hydrometallurgical conversion of chalcopyrite into low iron copper sulfides such as covellite or chalcocite. A summary of the processes based on hydrometallurgical conversion of chalcopyrite is described by Hackl, et al., in a paper entitled l'Reverse Leaching of Chalcopyrite," delivered at a meeting of the Canadian Institute of Mining in Toronto, Canada, in August, 1982. The pyrometallurgical processes involve roasting the concentrate to separate the undesired constituents as slag or volatilized off-gases. Such processes, however, require extensive environmental controls and yield waste products which are not readily disposed of.
There is also a significant loss of metal values to slag. In addition, the economics of scale are such that pyrometallurgical processes cannot usually be efficiently operated in association with single concentrators.
Consequently, it is not economically or environmentally feasible to improve low yrade concentrates through pyrometallurgy prior to shipping the concentrates for smelting .
Hydrometallurgical processes are based on the chemistry of reactions (1) and (2), which are illustrated with respect to the copper-iron-sulfur system:
CuFeS2 + Cu++ ~ 2CuS + Fe++ (1) FeS + Cu++ ~ CuS + Fe++ (2) These reactions are not limited to copper-iron-sulfur systems and analogous reactions can be written for other common metal sulfide systems, including nickel and zinc.
U.S. Patent 2,568~963 issued September 25, 19~7, to McGauley, et al., discloses a process for recovery of copper iZ~ 81 from an ore concentrate containing copper and iron sulfides.
The ore is treated with copper ions in solution and with a sulfuric acid-ferric sulfate solution in a substitution reaction to produce a copper sulfide containing slurry with a reduced iron sulfide content. The resulting slurry is filtered to produce a copper sulfide press cake which contains copper sulfide and a considerable amount of gangue (column 4, lines 66-68). The press cake is thereafter reacted with sulfuric acid to obtain copper sulfate which is treated with carbon monoxide to precipitate a high purity copper.
U.S. Patent 3,891,522 issued July 24, 1975, to McRay, et al., discloses a process for the production of elemental copper from copper-iron sulfides by reacting the copper~iron sulfides with metallic copper and a sulfuric acid solution to produce insoluble copper sulfide and a solution of ferrous sulfate; separating the copper sulfide from the ferrous sulfate and charging the copper sulfide as a slurry to an oxidizing leach where the copper sulfide is oxidized to solu~le copper sulfate for recovery as metallic copper. It is noted at column 10, line 46 to column 11, line 28 that quantities of gangues are associated with the copper sulfide and recovered as a residue from the oxidizing leach.
U.S. Patent 4,023,964 issued May 17, 1977, to DeMarthe, et al., was also considered in the preparation of this application.
Other processes are described in the Hackl, et al., reference and in Sohn, et al., J. Metals 18-22 (Nov , 1980) which use additional reductants along with a source of the 5~
desired metal. This is illustrated below in reactions (3), (4) and (5)/ again with respect to the copper-iron sulfur system:
CuFeS2 ~ Cu + 2H+ ~Cu2S + Fe++ + ~2S (3~
CuFeS2 + 3Cu++ + 2~2 -~ 2Cu2S + Fe++ + 4~+ (4) CuFeS2 + 3Cu++ + 2S~2 + 4H20 ~
2Cu2S + 6H+ + 2HSO4- + Fe++ (5) Although copper metal, hydrogen and SO2 are shown as reductants, otner reductants may be used.
It has been found that hydrometallurgical processes such as reactions (1) and (2) have not been effective to provide high enrichment of concentrates containing desired metal sulfides and mixed metal sulfides containing the desired metal and another metal and other metal sulfides for a variety of reasons. Attempts to completely convert all or a major portion of the undesired metals in the concentrates to the desired metal sulfides, which contain a high weight percent desired metal, by reacting the undesired metals with desired metal ions supplied in aqueous solution has resulted in difficulty in accomplishing complete conversion of the undesired metals and in excessive losses of desired metal upon separation of the aqueous solution containing desired metal ions. Further, most concentrates contain significant quantities of gangues which have been recovered with the minerals in such concentrates. Such gangues are still in the concentrates even after conversion of the minerals to desired metal sulfides and reduce the proportional concentration of the desired metal in the concentrates.
i2~8~81 Summary of the Invention The invention herein is a process for selectively enriching a sulfide mineral ore concentrate in a desired metal, the sulfide mineral ore concentrate being contained in an aqueous slurry of finely divided particulate solids containing desired metal sulfides, and mixed metal sulfides containing said desired metal and other metal or other metal sulfides, said other metal having a lower affinity for sulfur than the desired metal to produce an enriched concentrate having an increased desired metal content; said process comprising;
(a) reacting a desired metal ion rich aqueous stream with the mineral ore concentrate in a reaction zone to produce a product aqueous slurry containing increased quantities of desired metal sulfides;
(b) separating said product aqueous slurry into a desired metal ion depleted aqueous stream and a finely divided particulate solids stream containing the desired metal sulfides; and, (c) separating at least a portion of the desired metal sulfides from the finely divided particulate solids stream by a froth flotation process to produce the enriched concentrate.
The invention herein is particularly applicable to the enrichment of copper concentrates and could also be used to enrich concentrates of nickel, silver, gold, lead and zinc in which they are associated with elements such as iron.
This invention is particularly adapted to the recovery of an enriched copper concentrate from chalcopyrite ores.
~2~3i8~
The invention is also applicable to upgrading a copper sulfide (such as covellite) in copper by depleting the copper sulfide in sulfur.
Brief Description of Drawings Figure 1 is a schematic diagram of an embodiment of the present invention;
Figure 2 is a schematic diagram of a further embodiment of the present invention including a mechanical splitter to split the copper-iron concentrate into two streams;
Figure 3 is a schematic diagram of an embodiment of the present invention including a variation in the process of Figure 1 wherein selective flotation is used to control the portions of a copper-iron mineral ore concentrate passed to a dissolution zone and to a replacement reaction zone;
Figure ~ is a schematic diagram of an embodiment of the present invention wherein a selective flotation step is carried out in three stages for separation of the enriched concentrate;
Figure 5 is an idealized graph showing the composition of product solids and product solutions from an enrichment autoclave as a function of the ratio of copper in solution to the copper in the solids fed to the autoclave;
Figure 6 is a schematic diagram of a further embodiment of the present invention.
In the discussion of the Figures, the same numbers will be used throughout the discussion to refer to the same or similar components.
lZ513181 Description_of Preferred Embodiments In the Figures and in the discussions below, the invention will be discussed by reference to the enrichment of a copper concentrate in which the principal mineral is chalcopyrite although it is to be understood that the invention is not so limited, ~ut is also applicable to other physically and/or chemically mixed sulfides such as those containing copper, nickel, silver, gold, lead and zinc. The relationships between the different metals with respect to applicability of this process will be discussed below.
Throughout the specification, then, it will be recognized that while the principal discussion involves the upgrading of copper concentrates the invention has much wider applicability than for copper concentrates alone.
More specifically, the process of this invention is dependent on the relative solubilities of various metallic (or semimetallic) elements as sulfides at equilibrium, as set forth in the "Schurmann's Series" which is described in Park, et al., Ore DePosits (2nd Ed~, 1970), pages 478-484:
/ ~ palladium mercury silver bismuth increasing cadmium increasing affinity antimony solubility for sulfur tin as sulfide lead zinc nickel cobalt iron arsenic thallium manganese `
1258i~
In the present process, when an aqueous solution of a first metal (or metal ions) is placed in contact with solids containing a second metal which is lower in the Schurmann Series than the first metal, the first metal will displace the second metal from the solids. Thus, chalcopyrite (CuFeS2) is enriched in copper by displacement of iron when in contact with an aqueous stream containing copper ions.
It will be noted that the process of this invention can enhance the concentration of a desired metal in almost any concentrate (except those which are essentially at the maximum concentration). Of course the maximum benefit will be obtained when a "low grade" concentrate is enriched to a significantly higher grade. The invention is not limited, however, only to those concentrates which are commonly labeled "low grade" in the industry; rather the process is applicable to enhance the grade of any concentrate, even those which have an acceptable grade to start with.
An important advantage of this process is that a deeper cut can be made in the fixst selective concentration process (usually froth flotation) so that although a lower grade, i.e., lower weight percent desired metal concentrate is produced, the total desired metal recovery as a percent of the desired metal originally present in the ore is increased.
Previously it was necessary in some instances to accept a lower total desired metal recovery in order to achieve a higher grade concentrate. By the use of the present process, the higher total desired metal recovery can be achieved and the concentrate then upgraded so that both an enriched concentrate and a higher total desired metal recovery may be achieved.
~2S8i.81 Figure 5 illustrates the essential chemistry which is believed to underlie this invention~ When a copper sulfide mineral, such as chalcopyrite, is reacted in solution with a dissolved source of copper, under autoclave conditions of the type described in the prior art references, the conditions at any point are described by corresponding vertical points on lines I and II of the graph such as A/B and ~'/B'. Thus, a concentrate can be upgraded to Point B' in the autoclave but only at a cost of having a large amount of the copper remain in the autoclave solution, as indicated at A', and be lost to discharge unless recovered by another method. If the operator of the prior art processes wishes to minimize the amount of the desired metal which remains in solution (Point A), the operator must settle for a much lower degree of concentrate upgrading (Point B).
It is the discovery of our invention that by combining a replacement autoclave reaction and a subsequent control flotation of th~ autoclaved product in a manner not heretofore known or suggested by the prior art, we can achieve a high degree of concentrate upgrad.ing (Point B') while yet maintaining a minimum amount of loss of the copper (Point A).
While we do not wish to be bound by any specific theory, it is believed that the present process in effect operates such that the first stage of the upgrading (to Point B) is accomplished by the replacement reaction while the second stage of the upgrading (from Point B to Point B' or beyond) is accomplished by the control flotation~ By using the two steps in sequence, however, the loss of the desired metal which was a part of all prior art processes has been minimized such that the concentrate is upgraded from Point B to Point B' or beyond while the copper metal in solution remains at Point A. By the present process the concentrate may be enriched to a copper content approaching that for chalcocite. The enrichment accomplished by autoclaving, as discussed above, is subject to limitations on the degree of reaction completion and the presence of gangues or the other unreactive or unreacted solids in the concentrate, both of which limit the maximum enrichment which may be achieved. By the process of the lQ present invention such unreacted or unreactive materials are removed from the enriched concentrate to permit maximum enrichment of the concentrate.
Considering now this process with respect to Figure 1, a copper concentrate to be treated enters a reaction unit 8 (replacement reaction) through a line 7 and is dispersed in an aqueous medium containing copper ions. (As used herein, the term "aqueous medium" encompasses all water-based systems, including clear solutions, solids containing slurries and systems containing both dispersed and dissolved materials, as the context of the invention dictates.) If the concentrate were composed solely of chalcopyrite, the maximum copper content of the concentrate solids would be about 34 wt percent and typically about 25 to about 28 wt percent.
However, concentrates as noted normally contain a variety of other materials such as pyrite, chalcocite, gangue and other sulfides. Consequently, the effective proportion of copper in the incoming concentrate is normally about 25% and may be in the range of 10-35%. (All percentages herein are by weight unless otherwise stated.) i8~
The reactions in unit 8 are believed to proceed according to the following overall equation (~):
5CuFeS2 + llCU++ + 8~20 ~
8Cu2S + 5Fe++ ~ 2HSO4- + 14H~ (6) In the process of Figure 1 the source of the copper in solution is shown as cupric ion derived from an oxidation unit 42 (dissolution) and passed through a line 46 to unit 8.
The replacement reaction in unit 8 is normally carried on in an autoclave at elevated temperature and pressure, but as will be noted below, the use of additional reductants may allow for less severe reaction conditions.
Generally, however, it will be found that the operating temperature in reaction unit 8 will be in the range of about 80-350C (175-600F), preferably 120-300C (250-570F), and more preferably 160-230C (320-445F~. The reaction time will be at least 15 minutes, and preferably will be in the range of about 30-300 minutes, more preferably 60-120 minutes. The reaction temperature normally runs about 20C
(35F) below the reaction temperature in dissolution unit 42, since a proportion of the sensible heat generated in unit 42 and carried over with the solution containing the cupric ion is used to heat the incoming concentrate to the reaction temperature. Pressure in the autoclave (reaction unit 8) will be that pressure generated by the vapor pressure of the solution under the given temperature conditions. The reaction conditions stated above are those for chalcopyrite and may vary for other types of copper concentrates and other metalsO
The specific operating conditions for a given concentrate and given metal will be readily determined within known limits by those skilled in the art.
In some cases it may be desirable to add an additional reductant to unit 8 through a line 9. Such a reductant can be, e.g., metallic copper, sulfur, hydrogen or sulfur dioxide. These can be relatively cheap low-grade materials. Such addition will allow the operator to reduce the reaction temperature in unit 8. Such addition does, however, mean that there is an additional component in the system to be accounted for and separated. This is also another cost factor in the process which must be weighed against the economic savings from the reduced energy consumption.
It is desirable to control the acidity of this reaction by addition of a suitable neutralizing agent such as lime to prevent excessive solubilization of the iron oxides which may be carried into the reaction from the oxidation step. Desirably the pH in unit 8 is from about 0.8 to about
MINERAL ORE CONCENTXATES
Field of the Invention This invention relates to the selective enrichment of mineral ore concentratesn This invention further relates to the enrichment of mineral ore concentrates by a combination of a replacement step to produce additional quantities of sulfides of the desired metal in the concentrate solids and a selective froth flotation step to selectively recover the desired metal sulfides.
Description of the Prior Art In many commercial ore deposits, particularly those of copper, molybdenum, nickel, zinc and silver, the ore consists of a mixture of associated metal sulfide minerals.
Numerous minerals can be associated in any one orebody, and the specific minerals associated will vary from deposit to deposit. A large numbez of typical associations are described in Peters, Exploration and Mininq Geology (1978) Table 2-1.
For instance, in copper deposits in Utah and Chile, the copper-containlng minerals Chalcopyrite (CuFeS2), ~,r~
--1-- ~
bornite (cusFes4) and sometimes enargite (Cu3AsS4) are associated with pyrite (FeS2), molybdenite (MoS2) and in some cases sphalerite (ZnS) and galena (PbS). Similarly, a Montana copper deposit contains the copper-bearing minerals chalcopyrite, chalcocite (Cu2S), bornite, enargite, covellite (CuS) and sometimes tetrahedrite (CU3SbS3) and tennantite (Cu3Ass3) in association with pyrite and in some instances molybdenite, sphalerite and galena Such associations are not limited only to the copper deposits. Molybdenum sulfides are found in association with copper sulfides as noted above, and also with pyrite and magnetite (Fe3o4). Similarly, silver sulfides such as argentite (Ag2S) and polybasite (AgSbS6) are found in association with galena, chalcopyrite, pyrite and some tellurides. Nickel sulfides, such as pentlandite (Fe,Ni)gsg, are found in association with pyrrhotite (FeS), pyrite, chalcopyrite, marcasite (FeS2), magnetite and nickel arsenides. Zinc sulfides, such as sphalerite, are found associated with pyrrhotite, galena, chalcopyrite, pyrite and magnetite. In addition, it is common for some copper ores to contain small amounts of precious metal minera~s and trace metals such as bismuth, selenium and tellurium. It will be understood by those skilled in the art that the minerals listed above and mentioned hereinafter are described in idealized chemical form for convenience. It will be recognized, however, that most field ores actually have somewhat variable mineral composition and may include such minerals to a variable degree. See, for instance, E. S. Dana, A Textbook of Mineralogy, p 354 (4th ed., 1932).
1258~1 It will be noted from the preceding list that the deposits contain mixtures of both binary minerals such as chalcocite and covellite and ternary minerals such as chalcopyrite and bornite. More comple~ minerals may also be present.
In conventional mineral processing practice, it has been common to concentrate the ore containing metal values by processes such as grinding and flotation. The concentrate is then fed to a metal extraction process such as smelting to recover the desired metal values. The principal purpose of conventional ore concentration is two-fold: ~a) to separate minerals containing the desired metal from gangues (waste materials) to avoid having the gangues go through the desired metal separation process; and, (b) to separate those minerals containing the desired metal from minerals containing other desirable metals. Concentration is also beneficial when the concentrate of desired metal values is to be transported a substantial distance for smelting or other processing.
Increasing the weight percent desired metal in the concentrate reduces the weight and shipping costs of the concentrate per unit of desired metal in the concentrate.
The amount of upgrading possible by conventional concentration processes is limited by the chemical composition of the minerals containing the metal to be recovered. Further upgrading requires a change in the chemical composition of the minerals. There are two methods for enriching chalcopyrite copper concentrates by eliminating iron and sulfur. The first is conventional reverberatory roasting which eliminates iron in slag and sulfur as SO2. The second method involves 1~Z581~1 hydrometallurgical conversion of chalcopyrite into low iron copper sulfides such as covellite or chalcocite. A summary of the processes based on hydrometallurgical conversion of chalcopyrite is described by Hackl, et al., in a paper entitled l'Reverse Leaching of Chalcopyrite," delivered at a meeting of the Canadian Institute of Mining in Toronto, Canada, in August, 1982. The pyrometallurgical processes involve roasting the concentrate to separate the undesired constituents as slag or volatilized off-gases. Such processes, however, require extensive environmental controls and yield waste products which are not readily disposed of.
There is also a significant loss of metal values to slag. In addition, the economics of scale are such that pyrometallurgical processes cannot usually be efficiently operated in association with single concentrators.
Consequently, it is not economically or environmentally feasible to improve low yrade concentrates through pyrometallurgy prior to shipping the concentrates for smelting .
Hydrometallurgical processes are based on the chemistry of reactions (1) and (2), which are illustrated with respect to the copper-iron-sulfur system:
CuFeS2 + Cu++ ~ 2CuS + Fe++ (1) FeS + Cu++ ~ CuS + Fe++ (2) These reactions are not limited to copper-iron-sulfur systems and analogous reactions can be written for other common metal sulfide systems, including nickel and zinc.
U.S. Patent 2,568~963 issued September 25, 19~7, to McGauley, et al., discloses a process for recovery of copper iZ~ 81 from an ore concentrate containing copper and iron sulfides.
The ore is treated with copper ions in solution and with a sulfuric acid-ferric sulfate solution in a substitution reaction to produce a copper sulfide containing slurry with a reduced iron sulfide content. The resulting slurry is filtered to produce a copper sulfide press cake which contains copper sulfide and a considerable amount of gangue (column 4, lines 66-68). The press cake is thereafter reacted with sulfuric acid to obtain copper sulfate which is treated with carbon monoxide to precipitate a high purity copper.
U.S. Patent 3,891,522 issued July 24, 1975, to McRay, et al., discloses a process for the production of elemental copper from copper-iron sulfides by reacting the copper~iron sulfides with metallic copper and a sulfuric acid solution to produce insoluble copper sulfide and a solution of ferrous sulfate; separating the copper sulfide from the ferrous sulfate and charging the copper sulfide as a slurry to an oxidizing leach where the copper sulfide is oxidized to solu~le copper sulfate for recovery as metallic copper. It is noted at column 10, line 46 to column 11, line 28 that quantities of gangues are associated with the copper sulfide and recovered as a residue from the oxidizing leach.
U.S. Patent 4,023,964 issued May 17, 1977, to DeMarthe, et al., was also considered in the preparation of this application.
Other processes are described in the Hackl, et al., reference and in Sohn, et al., J. Metals 18-22 (Nov , 1980) which use additional reductants along with a source of the 5~
desired metal. This is illustrated below in reactions (3), (4) and (5)/ again with respect to the copper-iron sulfur system:
CuFeS2 ~ Cu + 2H+ ~Cu2S + Fe++ + ~2S (3~
CuFeS2 + 3Cu++ + 2~2 -~ 2Cu2S + Fe++ + 4~+ (4) CuFeS2 + 3Cu++ + 2S~2 + 4H20 ~
2Cu2S + 6H+ + 2HSO4- + Fe++ (5) Although copper metal, hydrogen and SO2 are shown as reductants, otner reductants may be used.
It has been found that hydrometallurgical processes such as reactions (1) and (2) have not been effective to provide high enrichment of concentrates containing desired metal sulfides and mixed metal sulfides containing the desired metal and another metal and other metal sulfides for a variety of reasons. Attempts to completely convert all or a major portion of the undesired metals in the concentrates to the desired metal sulfides, which contain a high weight percent desired metal, by reacting the undesired metals with desired metal ions supplied in aqueous solution has resulted in difficulty in accomplishing complete conversion of the undesired metals and in excessive losses of desired metal upon separation of the aqueous solution containing desired metal ions. Further, most concentrates contain significant quantities of gangues which have been recovered with the minerals in such concentrates. Such gangues are still in the concentrates even after conversion of the minerals to desired metal sulfides and reduce the proportional concentration of the desired metal in the concentrates.
i2~8~81 Summary of the Invention The invention herein is a process for selectively enriching a sulfide mineral ore concentrate in a desired metal, the sulfide mineral ore concentrate being contained in an aqueous slurry of finely divided particulate solids containing desired metal sulfides, and mixed metal sulfides containing said desired metal and other metal or other metal sulfides, said other metal having a lower affinity for sulfur than the desired metal to produce an enriched concentrate having an increased desired metal content; said process comprising;
(a) reacting a desired metal ion rich aqueous stream with the mineral ore concentrate in a reaction zone to produce a product aqueous slurry containing increased quantities of desired metal sulfides;
(b) separating said product aqueous slurry into a desired metal ion depleted aqueous stream and a finely divided particulate solids stream containing the desired metal sulfides; and, (c) separating at least a portion of the desired metal sulfides from the finely divided particulate solids stream by a froth flotation process to produce the enriched concentrate.
The invention herein is particularly applicable to the enrichment of copper concentrates and could also be used to enrich concentrates of nickel, silver, gold, lead and zinc in which they are associated with elements such as iron.
This invention is particularly adapted to the recovery of an enriched copper concentrate from chalcopyrite ores.
~2~3i8~
The invention is also applicable to upgrading a copper sulfide (such as covellite) in copper by depleting the copper sulfide in sulfur.
Brief Description of Drawings Figure 1 is a schematic diagram of an embodiment of the present invention;
Figure 2 is a schematic diagram of a further embodiment of the present invention including a mechanical splitter to split the copper-iron concentrate into two streams;
Figure 3 is a schematic diagram of an embodiment of the present invention including a variation in the process of Figure 1 wherein selective flotation is used to control the portions of a copper-iron mineral ore concentrate passed to a dissolution zone and to a replacement reaction zone;
Figure ~ is a schematic diagram of an embodiment of the present invention wherein a selective flotation step is carried out in three stages for separation of the enriched concentrate;
Figure 5 is an idealized graph showing the composition of product solids and product solutions from an enrichment autoclave as a function of the ratio of copper in solution to the copper in the solids fed to the autoclave;
Figure 6 is a schematic diagram of a further embodiment of the present invention.
In the discussion of the Figures, the same numbers will be used throughout the discussion to refer to the same or similar components.
lZ513181 Description_of Preferred Embodiments In the Figures and in the discussions below, the invention will be discussed by reference to the enrichment of a copper concentrate in which the principal mineral is chalcopyrite although it is to be understood that the invention is not so limited, ~ut is also applicable to other physically and/or chemically mixed sulfides such as those containing copper, nickel, silver, gold, lead and zinc. The relationships between the different metals with respect to applicability of this process will be discussed below.
Throughout the specification, then, it will be recognized that while the principal discussion involves the upgrading of copper concentrates the invention has much wider applicability than for copper concentrates alone.
More specifically, the process of this invention is dependent on the relative solubilities of various metallic (or semimetallic) elements as sulfides at equilibrium, as set forth in the "Schurmann's Series" which is described in Park, et al., Ore DePosits (2nd Ed~, 1970), pages 478-484:
/ ~ palladium mercury silver bismuth increasing cadmium increasing affinity antimony solubility for sulfur tin as sulfide lead zinc nickel cobalt iron arsenic thallium manganese `
1258i~
In the present process, when an aqueous solution of a first metal (or metal ions) is placed in contact with solids containing a second metal which is lower in the Schurmann Series than the first metal, the first metal will displace the second metal from the solids. Thus, chalcopyrite (CuFeS2) is enriched in copper by displacement of iron when in contact with an aqueous stream containing copper ions.
It will be noted that the process of this invention can enhance the concentration of a desired metal in almost any concentrate (except those which are essentially at the maximum concentration). Of course the maximum benefit will be obtained when a "low grade" concentrate is enriched to a significantly higher grade. The invention is not limited, however, only to those concentrates which are commonly labeled "low grade" in the industry; rather the process is applicable to enhance the grade of any concentrate, even those which have an acceptable grade to start with.
An important advantage of this process is that a deeper cut can be made in the fixst selective concentration process (usually froth flotation) so that although a lower grade, i.e., lower weight percent desired metal concentrate is produced, the total desired metal recovery as a percent of the desired metal originally present in the ore is increased.
Previously it was necessary in some instances to accept a lower total desired metal recovery in order to achieve a higher grade concentrate. By the use of the present process, the higher total desired metal recovery can be achieved and the concentrate then upgraded so that both an enriched concentrate and a higher total desired metal recovery may be achieved.
~2S8i.81 Figure 5 illustrates the essential chemistry which is believed to underlie this invention~ When a copper sulfide mineral, such as chalcopyrite, is reacted in solution with a dissolved source of copper, under autoclave conditions of the type described in the prior art references, the conditions at any point are described by corresponding vertical points on lines I and II of the graph such as A/B and ~'/B'. Thus, a concentrate can be upgraded to Point B' in the autoclave but only at a cost of having a large amount of the copper remain in the autoclave solution, as indicated at A', and be lost to discharge unless recovered by another method. If the operator of the prior art processes wishes to minimize the amount of the desired metal which remains in solution (Point A), the operator must settle for a much lower degree of concentrate upgrading (Point B).
It is the discovery of our invention that by combining a replacement autoclave reaction and a subsequent control flotation of th~ autoclaved product in a manner not heretofore known or suggested by the prior art, we can achieve a high degree of concentrate upgrad.ing (Point B') while yet maintaining a minimum amount of loss of the copper (Point A).
While we do not wish to be bound by any specific theory, it is believed that the present process in effect operates such that the first stage of the upgrading (to Point B) is accomplished by the replacement reaction while the second stage of the upgrading (from Point B to Point B' or beyond) is accomplished by the control flotation~ By using the two steps in sequence, however, the loss of the desired metal which was a part of all prior art processes has been minimized such that the concentrate is upgraded from Point B to Point B' or beyond while the copper metal in solution remains at Point A. By the present process the concentrate may be enriched to a copper content approaching that for chalcocite. The enrichment accomplished by autoclaving, as discussed above, is subject to limitations on the degree of reaction completion and the presence of gangues or the other unreactive or unreacted solids in the concentrate, both of which limit the maximum enrichment which may be achieved. By the process of the lQ present invention such unreacted or unreactive materials are removed from the enriched concentrate to permit maximum enrichment of the concentrate.
Considering now this process with respect to Figure 1, a copper concentrate to be treated enters a reaction unit 8 (replacement reaction) through a line 7 and is dispersed in an aqueous medium containing copper ions. (As used herein, the term "aqueous medium" encompasses all water-based systems, including clear solutions, solids containing slurries and systems containing both dispersed and dissolved materials, as the context of the invention dictates.) If the concentrate were composed solely of chalcopyrite, the maximum copper content of the concentrate solids would be about 34 wt percent and typically about 25 to about 28 wt percent.
However, concentrates as noted normally contain a variety of other materials such as pyrite, chalcocite, gangue and other sulfides. Consequently, the effective proportion of copper in the incoming concentrate is normally about 25% and may be in the range of 10-35%. (All percentages herein are by weight unless otherwise stated.) i8~
The reactions in unit 8 are believed to proceed according to the following overall equation (~):
5CuFeS2 + llCU++ + 8~20 ~
8Cu2S + 5Fe++ ~ 2HSO4- + 14H~ (6) In the process of Figure 1 the source of the copper in solution is shown as cupric ion derived from an oxidation unit 42 (dissolution) and passed through a line 46 to unit 8.
The replacement reaction in unit 8 is normally carried on in an autoclave at elevated temperature and pressure, but as will be noted below, the use of additional reductants may allow for less severe reaction conditions.
Generally, however, it will be found that the operating temperature in reaction unit 8 will be in the range of about 80-350C (175-600F), preferably 120-300C (250-570F), and more preferably 160-230C (320-445F~. The reaction time will be at least 15 minutes, and preferably will be in the range of about 30-300 minutes, more preferably 60-120 minutes. The reaction temperature normally runs about 20C
(35F) below the reaction temperature in dissolution unit 42, since a proportion of the sensible heat generated in unit 42 and carried over with the solution containing the cupric ion is used to heat the incoming concentrate to the reaction temperature. Pressure in the autoclave (reaction unit 8) will be that pressure generated by the vapor pressure of the solution under the given temperature conditions. The reaction conditions stated above are those for chalcopyrite and may vary for other types of copper concentrates and other metalsO
The specific operating conditions for a given concentrate and given metal will be readily determined within known limits by those skilled in the art.
In some cases it may be desirable to add an additional reductant to unit 8 through a line 9. Such a reductant can be, e.g., metallic copper, sulfur, hydrogen or sulfur dioxide. These can be relatively cheap low-grade materials. Such addition will allow the operator to reduce the reaction temperature in unit 8. Such addition does, however, mean that there is an additional component in the system to be accounted for and separated. This is also another cost factor in the process which must be weighed against the economic savings from the reduced energy consumption.
It is desirable to control the acidity of this reaction by addition of a suitable neutralizing agent such as lime to prevent excessive solubilization of the iron oxides which may be carried into the reaction from the oxidation step. Desirably the pH in unit 8 is from about 0.8 to about
2.2 and preferably from about 1.2 to about 1.6.
In the course of the reaction, the chalcopyrite is reacted in the presence of the cupric ion. The reaction as described above (reaction 6) results in the formation of chalcocite and/or covellite by the replacement of iron with copper and ferrous ion goes into solution. Sulfur may also be produced by conversion of chalcocite to covellite. The chalcocite and/or covellite becomes part of the concentrate, creating an "enriched" concentrate with respect to copper values. When the reaction has proceeded as far as desired by the operator, based on the chemical considerations illustrated 1'~58~
in Figure 4, the solids and solution are passed through a line 14 and subjected to conventional liquid/solid separation in a unit 16 such that solids are recovered and the ferrous ion and dissolved sulfur are discarded with the waste solution through a line 18. It would be possible to recycle a portion of the waste solution to dissolution unit 42 or unit 8 from unit 16, but such is not preferred since the waste solution contains substantial quantities of iron and possibly other undesirable materials. Unit 16 can also encompass drying of the solids to form a dry feed to be repulped for the subsequent control flotation. Desirably the reaction in unit 8 is controlled to react a selected portion of the copper-iron compounds in the concentrate stream to copper sulfides. The portion of the copper-iron compounds to be converted is dependent upon a variety of factors known to those skilled in the art such as the required residence time for the desired conversion, the ratio of copper to iron compounds initially present in the concentrate, process heat balance considerations, the amount of copper ions remaining in the discharged waste stream and the like. Desirably the amount of copper ions discharged in the waste stream is controlled to a value sufficiently low so that the waste stream can be discarded without further copper recovery therefrom. As indicated previously, it is not desirable that the waste stream be recycled. Copper could, of course, be recovered from this stream if desired. Desirably the copper concentration in the aqueous stream discharged through line 18 is less than about 1 gram per liter and preferably is less than about 0.5 grams per liter.
l'~S8~8:~
An important aspect of the present invention resides in the fact that the reaction just described is not used as the sole means of accomplishing the concentrate upgrading. In prior art processes, attempts were made to carry this reaction to completion with the result that there was loss of considerable copper ion to the waste solution. AS noted previously, even if the reaction is carried to completion, the presence of gangues and other unreactive or unreacted materials in the concentrate limits the maximum upgrading achievable. In the present process, however, the reaction is carried to an intermediate point (Point B in Figure 5) at which the operator determines that the loss of metal in solution (Point A in Figure 5) is at the optimum level from an economic and technical point of view. The solid material which is recovered from unit 16 through line 20 is ~herefore a "new" and enriched concentrate which contains not only the copper sulfides which have formed but also the unreacted portion of the incoming concentrate which may include pyrite, chalcocite, pyrite coated with chalcocite, unreacted chalcopyrite and gangue. These solid materials are passed to a control flotation unit 1~.
The control flotation process preerably consists of three different stages of selective flotation, subsequently producing three different copper concentrates, as illustrated in Figure 4.
In a first flotation stage 12A, a relatively high grade copper concentrate is produced. The concentrate consists primarily of copper sulfides and usually exceeds 50 percent copper. It will be understood that a variety of copper sulfide compounds are known and any or all of such copper sulfide compounds may be found in such recovered copper sulfides. Also reporting to this concentrate are mineral particles that are coated with chalcocite and consequently respond in the same manner as chalcocite. In this flotation stage 12A the uncoated pyrite is depressed, typically by using a source of sulfite ion such as SO2, alkali sulfite or bisulfite at a pulp pH of 2 to B. A copper selective collector (such as a thionocarbamate or dithiophosphate) is is then added in a limited amount and the more easily floated copper sulfides are collected. The pulp density is maintained between about 20 and about 40 percent solids and a frother is added as required to maintain adequate frothing conditions during flotation. This first concentrate is the product and is recovered through a line 22 for shipment as the enriched concentrate for further processing such as smelting.
In a second stage 12B of flotation, an additional amount of copper selective collector is added and an intermediate grade copper concentxate is produced. This concentrate usually exceeds 20 percent copper and contains chalcocite and covellite as the primary copper minerals.
These minerals may be physically attached to gangue minerals which dilutes the concentrate grade. Flotation is carried out in the same manner as was described in the initial flotation stage. The second concentrate makes a suitable feed for either oxidation dissolution unit 42 through a line 26 or enrichment stages in the enrichment process.
The significant difference between the two concentrates is that the first concentrate is higher grade and iZ~8~
contains the faster floating copper sulfides. On an operational basis, the first concentrate would correspond to the froth product recovered from the first set of flotation cells while the second concentrate would correspond to the froth product recovered from the second set of flotation cells.
In a third flotation stage 12C, the remaining sulfide minerals are collected from the gangue. This third concentrate usually exceeds two percent copper. It is collected from the cell product after the second flotation stage 12B by employing a collector (such as xanthate) that has a high degree of collecting power. With the exception of the collector, flotation is conducted in the same manner as was previously described. This third concentrate is subjected to a leach stage 32 to recover the copper while the cell product, containing mostly gangue, is rejected through a line 24.
Alternatively, gangue can be included with the third concentrate passing to the mild leach to dissolve copper-containing slimes in the gangue.
In the process of this invention, the control flotation step may be a single or multistage flotation. As described above, the preferred embodiment uses three stages.
~owever, both stages 12B and 12C are optional, and all the necessary cuts can, if desired, be made in stage 12Ao Alternatively, both 12A and 12B can be used, with 12C omitted.
Those skilled in the art will readily understand how the various cuts are made depending on the number of stages actually in use. Further, any of these concentrates can be subjected to additional flotation stages for further upgrading.
As noted, preferably the intermediate concentrate from flotation 12B is recycled through line 26 to dissolution step 42. It may be desirable to control the amount of concentrate recycled to dissolution step 42 so as to provide the total amount of copper for dissolution to produce the cupric ion needed in the process. This avoids the necessity of having to add any additional copper from outside sources.
Alternatively, as will be discussed below, the process of Figure 2 can be used to lessen the amount of recycle. The dissolution reaction is exothermic, so that commonly the process operates without the need for any additional energy input. In some cases, however, the heat generated by the dissolution will be insufficient to maintain the proper operating temperatures in units 42 and 8. Additional heat may be added to the system (by means not shown) through units 8 or ~2.
Typically, the process of this invention will upgrade a low grade concentrate in copper content by a factor of 2 to 3 times or greater. Iron can be reduced by a factor of up to 7 while sulfur is reduced by a factor of about two.
The amount of concentrate weight reduction can be approximately 2 to 3 times, so that the total quantity of concentrate to be shipped (usually to a smelter) is not only greatly enriched in copper and reduced in iron and sulfur, but also is a smaller total quantity, thus reducing the cost associated with shipping the concentrate.
The process as shown in Figure 1 includes a copper dissolution step in unit 42. The source for copper in solution may be an outside source, but it is believed that in 1~8~
most instances it will be preferred to produce the aqueous solution of copper ions in unit 42 using a portion of the concentrate or a process recycle stream. This step is sometimes referred to as an oxidation step because it can comprise a reaction to produce copper ions in solution from a copper source by oxidative reaction as well as by dissolution.
In preferred embodiments, the system will be as shown in Figures 1 and 2 in which the copper to be dissolved is provided at least in part by recycle within the system. Thus, a principal source of the copper to be oxidized is the intermediate recycle concentrate from line 26 which is obtained from stage 12B of flotation step 120 This recycle concentrate contains copper primarily in the form of chalcocite, covellite and chalcopyrite and is oxidized according to equations (7), (8), and (9) below:
2CuS + 4H+ + 502 ~~~~ 4Cu++ + 2S04= + 2H20 (7) CuS + 202 > Cu++ ~ SO4= (8) CuFeS2 ~ 402 --~ Cu++ + 2S04= + Fe++ (9) The copper bearing materials are dissolved in a conventional autoclave in which the temperature in unit 42 is maintained in the range oE at least 90C (195F), preferably in the range of 150-350C (300-660F), more preferably 180-250C (355-480F). Reaction time is at least about 15 minutes, and preferably in the range of 30 to 180 minutes.
As with all reaction conditions mentioned herein, these are subject to the known inverse time/temperature relationships such that increasing temperature requires less reaction time.
In addition, it will be understood that these specific conditions are applicable to the oxidation of copper in the 8~81 form of chalcocite and chalcopyri!e and that where other metals or copper in other forms are involved different but parallel reaction conditions will be used (some of which may overlap with stated reaction conditions). Finally, it will be also understood that these conditions may vary somewhat based on different types of concentrate, nature of the recycle stream and the presence of any foreign materials in the system which may have either a retarding or a catalytic effect on the oxidation. In all cases, however, it will be a routine matter for those skilled in the art to determine the optimum oxidation reaction conditions for the particular ore concentrates and metals that they are concerned with.
Normally oxygen is added to unit 42 through a line 44 by means such as gas sparging to bubble the oxygen through the copper solution. It is possible to add the oxygen through other forms of oxidants if desired; however, this will require that the other materials incorporated with the oxygen be accounted for and disposed of so that they will not accumulate in the system. It is desirable to control free acidity by maintaining the pH from about 0.8 to about 1.2 (as measured at ambient temperature) by the use of neutralizing agents such as lime. The oxidation reaction can be run over a wide pH range although the range stated is preferred for use in conjunction with the replacement reaction.
In a preferred embodiment, a portion of the third concentrate, and if desired gangue, is passed through a line 30 to an optional mild leach 32. The leach serves to remove dissolved chalcocite coatings on pyrite particles so that the dissolved copper can be recycled through a line 34 to 1~58~
oxidation and the pyrite discarded to waste through a line 36.
It also serves to dissolve very fine particles of copper minerals such as covellite and chalcocite which have escaped collection in the previous steps, so that ~he dissolved copper can be recycled. This leach will be particularly advantageous when the incoming low-grade concentrate contains a significant portion of chalcocite-coated pyrite, for such materials will pass virtually unchanged through the reaction in unit 8.
Flotation will not act to separate the rimmed chalcocite from the pyrite particles, even though the coating is only a small portion of the total particle weight, because the flotation process operates on the basis of surface chemistry and the chalcocite coating on the pyrite particles therefore reacts to flotation as if it were a particle solely of chalcocite. By leaching in unit 32 to dissolve the chalcocite coating, however, the pyrite can be readily separated and rejected to waste while the dissolved copper is recycled. A variety of well-known lixiviants, preferably mixtures of water with oxygen, air, ammonia or acids, can be used in leach 32 for the chalcocite recovery. In Figures 1 and 2 the process is exemplified by use of oxygen provided through line 33.
Conventional leaching operation conditions can be utilized.
In Figure 2 the copper-iron sulfide concentrate to be treated is charged through line 7 to a mechanical splitter zone 70 where the concentrate is split into two portions. A
first portion is passed through a line 72 to dissolution zone 42 with a second portion being passed through a line 74 to replacement reaction zone 8. The amount of concentrate charged to each zone will vary depending upon variables such as the amount of fresh concentrate required, in addition to the recycle concentrate from line 26 and copper solution from line 34, to produce the desired quantity of copper ions, the relative amounts of copper and iron in the concentrate, the process heat balance and the like~ 5uch variables are readily determined by those skilled in the art for any given concentrate. Generally, the use of a portion of the concentrate as a feed stream to dissolution zone 42 reduces the amount of relatively rich concentrate from second stage 12B flotation which must be recycled to dissolution vessel 42.
In general the operation of the process of Figure 2 is similar to operation of the process of Figure 1 except for the greater flexibility in the process disclosed in Figure 2.
In Figure 3, the copper-iron corcentrate to be treated is passed through line 7 to a selective flotation unit 70. The incoming concentrate contains, among other minerals, chalcocite and pyrite, and the purpose of selective flotation 70' is to effect a separation of the chalcocite from the pyrite. The pyrite-rich (copper leach) tail portion is passed through a line 74 directly to replacement reaction step 8. The chalcocite-rich froth (copper rich) floated portion is passed through a line 72 to dissolution (oxidation) step 42. This distribution of the pyrite can be used to control heat and acidity balances. This process variation is particularly desirable when the copper-iron concentrate contains high amounts of iron relative to the copper. In most instznces the oxidation (dissolution) of the copper mineral in dissolution zone 4~ is accompanied by some oxidation of pyrite. Since such oxidation reactions are exothermic, 12~3181 excessive amounts of heat may be generated in zone 42. Such poses a heat disposal problem in addition to requiring excessive amounts of oxygen. Further, the oxidation of such materials generates acid components and dissolved iron in the aqueous solution shown in reactions (7), (8) and (9) which requires the addition of neutralizing materials such as lime to maintain the desired pH range in zone 42 and in reaction zone 8. These problems are minimized by oxidizing the copper-rich portion of the copper-iron concentrate separated in selective flotation zone 70.
It should not noted that in the operation of this process embodiment, not only is all the chalcopyrite in the second portion of the concentrate charged to reaction zone 8 available for reaction with the copper ions, but the pyrite may also be reacted to produce copper sulfides. Variations in the extent of the flotation separation, the proportions of the copper-iron concentrate separated to the first portion and to the second portion and the process heat balance considerations to optimize the process operation are within the skill of those in the art based upon a given process configuration and a given concentrate. If chalcocite-rimmed pyrite is present, which is often the case, the chalcocite is separated from the pyrite by oxidation of the chalcocite in oxidation unit 42 in the process of Figure 3 to yield a pyrite product which may be rejected in subsequent processing.
In selective flotation unit 70', the concentrate is conditioned as a slurry with sulfuric acid until a pH of 2-8, preferably 5-6, is obtained. A source of sulfite ion such as SO2, allcali sulfite or bisulfite is then added (e.g., 1~8~81 approximately 1 9 of sodium sulfite per liter of slurry) and the pulp is further conditioned. The sulfite selectively depresses the pyrite so that it will not report to the froth (float) product. A small amount (e.g., 0.1 lb/ton) of a copper selective collector tsuch as a thionocarbamate or dithiophosphate) is then added and the pulp subjected to the flotation. Frother is added as required to maintain adequate frothing conditions. The chalcocite reports to the floated product while the pyrite remains in the flotation cell as part of the tail. Other copper materials which may be present, such as bornite, chalcopyrite and covellite split approximately evenly between the float and the tail.
In replacement reaction unit 8 the tail portion of the concentrate is dispersed in an aqueous medium containing the dissolved source of the cupric ions and the operation of the process in Figure 3 is otherwise generally the same as discussed in conjunction with Figure 1.
In a variation of the process disclosed herein an ammonia oxidation may be used in dissolution unit 42. The equipment and operating conditions in unit 42 are substantially the same as used previously except that the pH
in the oxidation reaction is maintained at a value from about 9.5 to about 10.5 by the addition of ammonia and lime (not shown). The oxidation reaction in the ammonia system may be illustrated as in equation (10) below.
5/202 + Cu2S + H20 + 4NH3 2Cu(NH3)4~+ + S04= + 2(0H)- (10) The main reactions for the production of copper ions in the oxidation autoclave of an acid circuit have been shown ~;~S~1~8~ -previously. Chalcopyrite does not oxidize rapidly in an ammonia circuit because an insoluble iron oxide layer forms on the chalcopyrite grains.
The reaction in reactor 8 with the copper-ammonia ions is virtually identical with the reaction in reactor 8 for the acid circuit except the reacting copper species is Cu(N~3)4++ rather than Cu++. The difference in reaction products and the low solubility of iron in the basic circuit will change the heat balance slightly, but since iron is not soluble in the basic circuit it may be possible to operate at higher copper ion concentration.
The ammonia system results in a much lower acid production, and is better for copper-iron concentrates having a high pyrite content. The ammonia system would be very effective in the treatment of chalcocite-rimmed pyrite.
In a further variation of the present process as shown in Figure 6, all the concentrate is charged to a single multistage autoclave 100. While not discussed previously, it should be understood that the equipment used for both the dissolution and the replacement reaction steps may be and preferably is multistage autoclaves. In a multistage autoclave 100 shown in Figure 7 which contains stages 100a-h, the copper-iron concentrate is charged to a first stage 100a and then sequentially passes to and through subsequent stages 100b-h to discharge through a line 14. Oxygen or other suitable free-oxygen containing gas such as air, oxygen enriched air or the like is injected through a line 44 to first stage 100a oxidize and solubilize a selected portion of the copper present in the concentrate. First stage 100a may .~Z58~8~
be larger than the subsequent stages and desirably most of the oxygen and lime will be added to first stage 100a. Generally from about 30 to about 70 percent of the copper in the concentrate and from about 20 to about 50 percent of the iron in the concentrate may be oxidized. Optionally recycled desired metal containing streams such as from lines 26 and 34 may also be charged to autoclave 100. Oxygen may be added to subsequent stages as shown by line 44' although it is desirable that most oxidation occur in the first stage or at least the first few stages. The conditions in the dissolution zone (corresponding to zone 42 in Figure 1 and shown as stages 100a-c in Figure 7) are generally the same as in zone 42 discussed in conjunction with Figure 1. Similarly, conditions in the replace~ent reaction zone (corresponding to reaction zone 8 in Figure 1 and shown as stages 100d-h) are generally the same as in reactive zone 8. Desirably the pH is controlled in autoclave 100 by lime addition or the like (not shown) as required to maintain the pH in the desired range as discussed previously for the oxidation and reaction zones to prevent precipitation of iron. The product aqueous slurry recovered via line 14 is passed to further processing as discussed previously.
In the process disclosed above, the precious metals (such as silver, gold and platinum group metals) which are frequently associated with copper in sulfides are effectively carried along with the copper and appear in the upgraded concentrate, themselves in upgraded content as compared to their content in the initial concentrate.
12~
The p~ocess also serves to reduce the content of trace elements such as bismuth, arsenic, selenium and tellurium in the enriched concentrate. Such elements are hard to remove by conventional beneficiation techniques. The reduction of these elements has at least two benefits: it improves the purity of the subsequent copper products and it reduces the cost of smeltiny, for smelters usually charge "penalty fees" for handling concentrates which contain various quantities of these undesirable materials.
In the present process, since it is not necessary to convert the entire solids content of the concentrate to the desired metal sulfides, it is not necessary to attempt a complete reaction by maintaining a high desired metal ion content in the aqueous stream in contact with the solids throughout the replacement reaction. As a result, the aqueous stream in contact with the solids can be allowed to drop to a relatively low desired metal ion content at the end of the reaction time. As a result, the waste aqueous solution discarded from the replacement reaction can be relatively low in the desired metal ion as discussed previously.
Since the replacement reaction is not the only step relied upon for the enrichment of the concentrate, it is not necessary that this reaction be carried to completion. As discussed previously, even were it attempted to carry this reaction to completion certain unreacted materials would remain with the concentrate as well as gangues and the like which were initially present in the concentrate materials as a result of their recovery in the previous concentration step.
Such materials tend to reduce the maximum enrichment possible in the replacement reaction~
:~S8~8~
In the present process the enrichment is accomplished by two steps in combination which synergistically cooperate to produce a very high degree of desired metal enrichment in the product concentrate. The second step as discussed previously is a flotation step where the enriched materials, i.e. primarily desired metal sulfides, are selectively separated from the less enriched desired metal con~aining compounds remaining in the concentrate and from the gangues and other unreacted or unreactive materials. As indicated previously, these two steps, in combination synergistically, cooperate to produce an enriched concentrate efficiently. In other words, it is not necessary to attempt to drive the replacement reaction to completion since a means is available for selectively recovering only the enriched product produced. The unreacted or unreactive materials can conveniently be separated for recycle or other uses in the process or discharged to waste. As a result, the waste stream discarded from the replacement reaction which contains undesirable iron and other materials may contain minimal amounts of the desired metal ions. Similarly the use of the control flotation step results in the ability to recover the most desirable materials from the resultant product from the replacement reaction without the necessity for discarding valuable products since the recovery of these products in second and third stage flotation steps permits their recovery and recycle to beneficial use in the process.
Having thus described the invention by reference to certain of its preferred embodiments, it is respectfully pointed out that the embodiments described are illustrative :12S~3~81 rather than limiting in nature and that many variations and modifications are possible within the scope of the present invention. Many such variations and modifications may be considered obvious and desirable by those skilled in the art based upon a review of the foregoing description of preferred embodiments.
Having thus described the invention, we claim:
In the course of the reaction, the chalcopyrite is reacted in the presence of the cupric ion. The reaction as described above (reaction 6) results in the formation of chalcocite and/or covellite by the replacement of iron with copper and ferrous ion goes into solution. Sulfur may also be produced by conversion of chalcocite to covellite. The chalcocite and/or covellite becomes part of the concentrate, creating an "enriched" concentrate with respect to copper values. When the reaction has proceeded as far as desired by the operator, based on the chemical considerations illustrated 1'~58~
in Figure 4, the solids and solution are passed through a line 14 and subjected to conventional liquid/solid separation in a unit 16 such that solids are recovered and the ferrous ion and dissolved sulfur are discarded with the waste solution through a line 18. It would be possible to recycle a portion of the waste solution to dissolution unit 42 or unit 8 from unit 16, but such is not preferred since the waste solution contains substantial quantities of iron and possibly other undesirable materials. Unit 16 can also encompass drying of the solids to form a dry feed to be repulped for the subsequent control flotation. Desirably the reaction in unit 8 is controlled to react a selected portion of the copper-iron compounds in the concentrate stream to copper sulfides. The portion of the copper-iron compounds to be converted is dependent upon a variety of factors known to those skilled in the art such as the required residence time for the desired conversion, the ratio of copper to iron compounds initially present in the concentrate, process heat balance considerations, the amount of copper ions remaining in the discharged waste stream and the like. Desirably the amount of copper ions discharged in the waste stream is controlled to a value sufficiently low so that the waste stream can be discarded without further copper recovery therefrom. As indicated previously, it is not desirable that the waste stream be recycled. Copper could, of course, be recovered from this stream if desired. Desirably the copper concentration in the aqueous stream discharged through line 18 is less than about 1 gram per liter and preferably is less than about 0.5 grams per liter.
l'~S8~8:~
An important aspect of the present invention resides in the fact that the reaction just described is not used as the sole means of accomplishing the concentrate upgrading. In prior art processes, attempts were made to carry this reaction to completion with the result that there was loss of considerable copper ion to the waste solution. AS noted previously, even if the reaction is carried to completion, the presence of gangues and other unreactive or unreacted materials in the concentrate limits the maximum upgrading achievable. In the present process, however, the reaction is carried to an intermediate point (Point B in Figure 5) at which the operator determines that the loss of metal in solution (Point A in Figure 5) is at the optimum level from an economic and technical point of view. The solid material which is recovered from unit 16 through line 20 is ~herefore a "new" and enriched concentrate which contains not only the copper sulfides which have formed but also the unreacted portion of the incoming concentrate which may include pyrite, chalcocite, pyrite coated with chalcocite, unreacted chalcopyrite and gangue. These solid materials are passed to a control flotation unit 1~.
The control flotation process preerably consists of three different stages of selective flotation, subsequently producing three different copper concentrates, as illustrated in Figure 4.
In a first flotation stage 12A, a relatively high grade copper concentrate is produced. The concentrate consists primarily of copper sulfides and usually exceeds 50 percent copper. It will be understood that a variety of copper sulfide compounds are known and any or all of such copper sulfide compounds may be found in such recovered copper sulfides. Also reporting to this concentrate are mineral particles that are coated with chalcocite and consequently respond in the same manner as chalcocite. In this flotation stage 12A the uncoated pyrite is depressed, typically by using a source of sulfite ion such as SO2, alkali sulfite or bisulfite at a pulp pH of 2 to B. A copper selective collector (such as a thionocarbamate or dithiophosphate) is is then added in a limited amount and the more easily floated copper sulfides are collected. The pulp density is maintained between about 20 and about 40 percent solids and a frother is added as required to maintain adequate frothing conditions during flotation. This first concentrate is the product and is recovered through a line 22 for shipment as the enriched concentrate for further processing such as smelting.
In a second stage 12B of flotation, an additional amount of copper selective collector is added and an intermediate grade copper concentxate is produced. This concentrate usually exceeds 20 percent copper and contains chalcocite and covellite as the primary copper minerals.
These minerals may be physically attached to gangue minerals which dilutes the concentrate grade. Flotation is carried out in the same manner as was described in the initial flotation stage. The second concentrate makes a suitable feed for either oxidation dissolution unit 42 through a line 26 or enrichment stages in the enrichment process.
The significant difference between the two concentrates is that the first concentrate is higher grade and iZ~8~
contains the faster floating copper sulfides. On an operational basis, the first concentrate would correspond to the froth product recovered from the first set of flotation cells while the second concentrate would correspond to the froth product recovered from the second set of flotation cells.
In a third flotation stage 12C, the remaining sulfide minerals are collected from the gangue. This third concentrate usually exceeds two percent copper. It is collected from the cell product after the second flotation stage 12B by employing a collector (such as xanthate) that has a high degree of collecting power. With the exception of the collector, flotation is conducted in the same manner as was previously described. This third concentrate is subjected to a leach stage 32 to recover the copper while the cell product, containing mostly gangue, is rejected through a line 24.
Alternatively, gangue can be included with the third concentrate passing to the mild leach to dissolve copper-containing slimes in the gangue.
In the process of this invention, the control flotation step may be a single or multistage flotation. As described above, the preferred embodiment uses three stages.
~owever, both stages 12B and 12C are optional, and all the necessary cuts can, if desired, be made in stage 12Ao Alternatively, both 12A and 12B can be used, with 12C omitted.
Those skilled in the art will readily understand how the various cuts are made depending on the number of stages actually in use. Further, any of these concentrates can be subjected to additional flotation stages for further upgrading.
As noted, preferably the intermediate concentrate from flotation 12B is recycled through line 26 to dissolution step 42. It may be desirable to control the amount of concentrate recycled to dissolution step 42 so as to provide the total amount of copper for dissolution to produce the cupric ion needed in the process. This avoids the necessity of having to add any additional copper from outside sources.
Alternatively, as will be discussed below, the process of Figure 2 can be used to lessen the amount of recycle. The dissolution reaction is exothermic, so that commonly the process operates without the need for any additional energy input. In some cases, however, the heat generated by the dissolution will be insufficient to maintain the proper operating temperatures in units 42 and 8. Additional heat may be added to the system (by means not shown) through units 8 or ~2.
Typically, the process of this invention will upgrade a low grade concentrate in copper content by a factor of 2 to 3 times or greater. Iron can be reduced by a factor of up to 7 while sulfur is reduced by a factor of about two.
The amount of concentrate weight reduction can be approximately 2 to 3 times, so that the total quantity of concentrate to be shipped (usually to a smelter) is not only greatly enriched in copper and reduced in iron and sulfur, but also is a smaller total quantity, thus reducing the cost associated with shipping the concentrate.
The process as shown in Figure 1 includes a copper dissolution step in unit 42. The source for copper in solution may be an outside source, but it is believed that in 1~8~
most instances it will be preferred to produce the aqueous solution of copper ions in unit 42 using a portion of the concentrate or a process recycle stream. This step is sometimes referred to as an oxidation step because it can comprise a reaction to produce copper ions in solution from a copper source by oxidative reaction as well as by dissolution.
In preferred embodiments, the system will be as shown in Figures 1 and 2 in which the copper to be dissolved is provided at least in part by recycle within the system. Thus, a principal source of the copper to be oxidized is the intermediate recycle concentrate from line 26 which is obtained from stage 12B of flotation step 120 This recycle concentrate contains copper primarily in the form of chalcocite, covellite and chalcopyrite and is oxidized according to equations (7), (8), and (9) below:
2CuS + 4H+ + 502 ~~~~ 4Cu++ + 2S04= + 2H20 (7) CuS + 202 > Cu++ ~ SO4= (8) CuFeS2 ~ 402 --~ Cu++ + 2S04= + Fe++ (9) The copper bearing materials are dissolved in a conventional autoclave in which the temperature in unit 42 is maintained in the range oE at least 90C (195F), preferably in the range of 150-350C (300-660F), more preferably 180-250C (355-480F). Reaction time is at least about 15 minutes, and preferably in the range of 30 to 180 minutes.
As with all reaction conditions mentioned herein, these are subject to the known inverse time/temperature relationships such that increasing temperature requires less reaction time.
In addition, it will be understood that these specific conditions are applicable to the oxidation of copper in the 8~81 form of chalcocite and chalcopyri!e and that where other metals or copper in other forms are involved different but parallel reaction conditions will be used (some of which may overlap with stated reaction conditions). Finally, it will be also understood that these conditions may vary somewhat based on different types of concentrate, nature of the recycle stream and the presence of any foreign materials in the system which may have either a retarding or a catalytic effect on the oxidation. In all cases, however, it will be a routine matter for those skilled in the art to determine the optimum oxidation reaction conditions for the particular ore concentrates and metals that they are concerned with.
Normally oxygen is added to unit 42 through a line 44 by means such as gas sparging to bubble the oxygen through the copper solution. It is possible to add the oxygen through other forms of oxidants if desired; however, this will require that the other materials incorporated with the oxygen be accounted for and disposed of so that they will not accumulate in the system. It is desirable to control free acidity by maintaining the pH from about 0.8 to about 1.2 (as measured at ambient temperature) by the use of neutralizing agents such as lime. The oxidation reaction can be run over a wide pH range although the range stated is preferred for use in conjunction with the replacement reaction.
In a preferred embodiment, a portion of the third concentrate, and if desired gangue, is passed through a line 30 to an optional mild leach 32. The leach serves to remove dissolved chalcocite coatings on pyrite particles so that the dissolved copper can be recycled through a line 34 to 1~58~
oxidation and the pyrite discarded to waste through a line 36.
It also serves to dissolve very fine particles of copper minerals such as covellite and chalcocite which have escaped collection in the previous steps, so that ~he dissolved copper can be recycled. This leach will be particularly advantageous when the incoming low-grade concentrate contains a significant portion of chalcocite-coated pyrite, for such materials will pass virtually unchanged through the reaction in unit 8.
Flotation will not act to separate the rimmed chalcocite from the pyrite particles, even though the coating is only a small portion of the total particle weight, because the flotation process operates on the basis of surface chemistry and the chalcocite coating on the pyrite particles therefore reacts to flotation as if it were a particle solely of chalcocite. By leaching in unit 32 to dissolve the chalcocite coating, however, the pyrite can be readily separated and rejected to waste while the dissolved copper is recycled. A variety of well-known lixiviants, preferably mixtures of water with oxygen, air, ammonia or acids, can be used in leach 32 for the chalcocite recovery. In Figures 1 and 2 the process is exemplified by use of oxygen provided through line 33.
Conventional leaching operation conditions can be utilized.
In Figure 2 the copper-iron sulfide concentrate to be treated is charged through line 7 to a mechanical splitter zone 70 where the concentrate is split into two portions. A
first portion is passed through a line 72 to dissolution zone 42 with a second portion being passed through a line 74 to replacement reaction zone 8. The amount of concentrate charged to each zone will vary depending upon variables such as the amount of fresh concentrate required, in addition to the recycle concentrate from line 26 and copper solution from line 34, to produce the desired quantity of copper ions, the relative amounts of copper and iron in the concentrate, the process heat balance and the like~ 5uch variables are readily determined by those skilled in the art for any given concentrate. Generally, the use of a portion of the concentrate as a feed stream to dissolution zone 42 reduces the amount of relatively rich concentrate from second stage 12B flotation which must be recycled to dissolution vessel 42.
In general the operation of the process of Figure 2 is similar to operation of the process of Figure 1 except for the greater flexibility in the process disclosed in Figure 2.
In Figure 3, the copper-iron corcentrate to be treated is passed through line 7 to a selective flotation unit 70. The incoming concentrate contains, among other minerals, chalcocite and pyrite, and the purpose of selective flotation 70' is to effect a separation of the chalcocite from the pyrite. The pyrite-rich (copper leach) tail portion is passed through a line 74 directly to replacement reaction step 8. The chalcocite-rich froth (copper rich) floated portion is passed through a line 72 to dissolution (oxidation) step 42. This distribution of the pyrite can be used to control heat and acidity balances. This process variation is particularly desirable when the copper-iron concentrate contains high amounts of iron relative to the copper. In most instznces the oxidation (dissolution) of the copper mineral in dissolution zone 4~ is accompanied by some oxidation of pyrite. Since such oxidation reactions are exothermic, 12~3181 excessive amounts of heat may be generated in zone 42. Such poses a heat disposal problem in addition to requiring excessive amounts of oxygen. Further, the oxidation of such materials generates acid components and dissolved iron in the aqueous solution shown in reactions (7), (8) and (9) which requires the addition of neutralizing materials such as lime to maintain the desired pH range in zone 42 and in reaction zone 8. These problems are minimized by oxidizing the copper-rich portion of the copper-iron concentrate separated in selective flotation zone 70.
It should not noted that in the operation of this process embodiment, not only is all the chalcopyrite in the second portion of the concentrate charged to reaction zone 8 available for reaction with the copper ions, but the pyrite may also be reacted to produce copper sulfides. Variations in the extent of the flotation separation, the proportions of the copper-iron concentrate separated to the first portion and to the second portion and the process heat balance considerations to optimize the process operation are within the skill of those in the art based upon a given process configuration and a given concentrate. If chalcocite-rimmed pyrite is present, which is often the case, the chalcocite is separated from the pyrite by oxidation of the chalcocite in oxidation unit 42 in the process of Figure 3 to yield a pyrite product which may be rejected in subsequent processing.
In selective flotation unit 70', the concentrate is conditioned as a slurry with sulfuric acid until a pH of 2-8, preferably 5-6, is obtained. A source of sulfite ion such as SO2, allcali sulfite or bisulfite is then added (e.g., 1~8~81 approximately 1 9 of sodium sulfite per liter of slurry) and the pulp is further conditioned. The sulfite selectively depresses the pyrite so that it will not report to the froth (float) product. A small amount (e.g., 0.1 lb/ton) of a copper selective collector tsuch as a thionocarbamate or dithiophosphate) is then added and the pulp subjected to the flotation. Frother is added as required to maintain adequate frothing conditions. The chalcocite reports to the floated product while the pyrite remains in the flotation cell as part of the tail. Other copper materials which may be present, such as bornite, chalcopyrite and covellite split approximately evenly between the float and the tail.
In replacement reaction unit 8 the tail portion of the concentrate is dispersed in an aqueous medium containing the dissolved source of the cupric ions and the operation of the process in Figure 3 is otherwise generally the same as discussed in conjunction with Figure 1.
In a variation of the process disclosed herein an ammonia oxidation may be used in dissolution unit 42. The equipment and operating conditions in unit 42 are substantially the same as used previously except that the pH
in the oxidation reaction is maintained at a value from about 9.5 to about 10.5 by the addition of ammonia and lime (not shown). The oxidation reaction in the ammonia system may be illustrated as in equation (10) below.
5/202 + Cu2S + H20 + 4NH3 2Cu(NH3)4~+ + S04= + 2(0H)- (10) The main reactions for the production of copper ions in the oxidation autoclave of an acid circuit have been shown ~;~S~1~8~ -previously. Chalcopyrite does not oxidize rapidly in an ammonia circuit because an insoluble iron oxide layer forms on the chalcopyrite grains.
The reaction in reactor 8 with the copper-ammonia ions is virtually identical with the reaction in reactor 8 for the acid circuit except the reacting copper species is Cu(N~3)4++ rather than Cu++. The difference in reaction products and the low solubility of iron in the basic circuit will change the heat balance slightly, but since iron is not soluble in the basic circuit it may be possible to operate at higher copper ion concentration.
The ammonia system results in a much lower acid production, and is better for copper-iron concentrates having a high pyrite content. The ammonia system would be very effective in the treatment of chalcocite-rimmed pyrite.
In a further variation of the present process as shown in Figure 6, all the concentrate is charged to a single multistage autoclave 100. While not discussed previously, it should be understood that the equipment used for both the dissolution and the replacement reaction steps may be and preferably is multistage autoclaves. In a multistage autoclave 100 shown in Figure 7 which contains stages 100a-h, the copper-iron concentrate is charged to a first stage 100a and then sequentially passes to and through subsequent stages 100b-h to discharge through a line 14. Oxygen or other suitable free-oxygen containing gas such as air, oxygen enriched air or the like is injected through a line 44 to first stage 100a oxidize and solubilize a selected portion of the copper present in the concentrate. First stage 100a may .~Z58~8~
be larger than the subsequent stages and desirably most of the oxygen and lime will be added to first stage 100a. Generally from about 30 to about 70 percent of the copper in the concentrate and from about 20 to about 50 percent of the iron in the concentrate may be oxidized. Optionally recycled desired metal containing streams such as from lines 26 and 34 may also be charged to autoclave 100. Oxygen may be added to subsequent stages as shown by line 44' although it is desirable that most oxidation occur in the first stage or at least the first few stages. The conditions in the dissolution zone (corresponding to zone 42 in Figure 1 and shown as stages 100a-c in Figure 7) are generally the same as in zone 42 discussed in conjunction with Figure 1. Similarly, conditions in the replace~ent reaction zone (corresponding to reaction zone 8 in Figure 1 and shown as stages 100d-h) are generally the same as in reactive zone 8. Desirably the pH is controlled in autoclave 100 by lime addition or the like (not shown) as required to maintain the pH in the desired range as discussed previously for the oxidation and reaction zones to prevent precipitation of iron. The product aqueous slurry recovered via line 14 is passed to further processing as discussed previously.
In the process disclosed above, the precious metals (such as silver, gold and platinum group metals) which are frequently associated with copper in sulfides are effectively carried along with the copper and appear in the upgraded concentrate, themselves in upgraded content as compared to their content in the initial concentrate.
12~
The p~ocess also serves to reduce the content of trace elements such as bismuth, arsenic, selenium and tellurium in the enriched concentrate. Such elements are hard to remove by conventional beneficiation techniques. The reduction of these elements has at least two benefits: it improves the purity of the subsequent copper products and it reduces the cost of smeltiny, for smelters usually charge "penalty fees" for handling concentrates which contain various quantities of these undesirable materials.
In the present process, since it is not necessary to convert the entire solids content of the concentrate to the desired metal sulfides, it is not necessary to attempt a complete reaction by maintaining a high desired metal ion content in the aqueous stream in contact with the solids throughout the replacement reaction. As a result, the aqueous stream in contact with the solids can be allowed to drop to a relatively low desired metal ion content at the end of the reaction time. As a result, the waste aqueous solution discarded from the replacement reaction can be relatively low in the desired metal ion as discussed previously.
Since the replacement reaction is not the only step relied upon for the enrichment of the concentrate, it is not necessary that this reaction be carried to completion. As discussed previously, even were it attempted to carry this reaction to completion certain unreacted materials would remain with the concentrate as well as gangues and the like which were initially present in the concentrate materials as a result of their recovery in the previous concentration step.
Such materials tend to reduce the maximum enrichment possible in the replacement reaction~
:~S8~8~
In the present process the enrichment is accomplished by two steps in combination which synergistically cooperate to produce a very high degree of desired metal enrichment in the product concentrate. The second step as discussed previously is a flotation step where the enriched materials, i.e. primarily desired metal sulfides, are selectively separated from the less enriched desired metal con~aining compounds remaining in the concentrate and from the gangues and other unreacted or unreactive materials. As indicated previously, these two steps, in combination synergistically, cooperate to produce an enriched concentrate efficiently. In other words, it is not necessary to attempt to drive the replacement reaction to completion since a means is available for selectively recovering only the enriched product produced. The unreacted or unreactive materials can conveniently be separated for recycle or other uses in the process or discharged to waste. As a result, the waste stream discarded from the replacement reaction which contains undesirable iron and other materials may contain minimal amounts of the desired metal ions. Similarly the use of the control flotation step results in the ability to recover the most desirable materials from the resultant product from the replacement reaction without the necessity for discarding valuable products since the recovery of these products in second and third stage flotation steps permits their recovery and recycle to beneficial use in the process.
Having thus described the invention by reference to certain of its preferred embodiments, it is respectfully pointed out that the embodiments described are illustrative :12S~3~81 rather than limiting in nature and that many variations and modifications are possible within the scope of the present invention. Many such variations and modifications may be considered obvious and desirable by those skilled in the art based upon a review of the foregoing description of preferred embodiments.
Having thus described the invention, we claim:
Claims (35)
1. A process for selectively enriching a sulfide mineral ore concentrate in a desired metal, said sulfide mineral ore concentrate being contained in an aqueous slurry of finely divided particulate solids containing desired metal sulfides and mixed metal sulfides containing said desired metal and other metal or other metal sulfides, said other metal having a lower affinity for sulfur than said desired metal to produce an enriched concentrate having an increased desired metal content; said process comprising:
(a) reacting a desired metal ion rich aqueous stream with said sulfide mineral ore concentrate in a reaction zone to produce a product aqueous slurry containing increased quantities of desired metal sulfides;
(b) separating said product aqueous slurry into a desired metal ion depleted aqueous stream and a finely divided particulate solids stream containing said desired metal sulfides; and, (c) separating at least a portion of said desired metal sulfides from said finely divided particulate solids stream by a froth flotation process to produce said enriched concentrate.
(a) reacting a desired metal ion rich aqueous stream with said sulfide mineral ore concentrate in a reaction zone to produce a product aqueous slurry containing increased quantities of desired metal sulfides;
(b) separating said product aqueous slurry into a desired metal ion depleted aqueous stream and a finely divided particulate solids stream containing said desired metal sulfides; and, (c) separating at least a portion of said desired metal sulfides from said finely divided particulate solids stream by a froth flotation process to produce said enriched concentrate.
2. The process of Claim 1 wherein said desired metal is selected from the group consisting of copper, nickel, silver, gold, lead and zinc.
3. The process of Claim 2 wherein said enriched concentrate contains more than about 50 weight percent desired metal.
4. The process of Claim 2 wherein said desired metal is copper.
5. The process of Claim 4 wherein said enriched concentrate contains more than about 50 weight percent copper.
6. The process of Claim 4 wherein said mineral ore concentrate contains copper-iron sulfides.
7. The process of Claim 4 wherein said reaction is at a temperature from about 80 to about 350°C for a time greater than about 15 minutes.
8. The process of Claim 7 wherein said reaction is at a temperature from about 120 to about 300°C for a time from about 30 to about 300 minutes.
9. The process of Claim 4 wherein said copper ion depleted aqueous stream contains less than about 1 gram per liter of dissolved copper.
10. The process of Claim 9 wherein said copper ion depleted aqueous stream contains less than about 0.5 grams per liter of dissolved copper.
11. The process of Claim 1 wherein said froth flotation process comprises a plurality of individual flotation steps in series, at least the first of which separates said desired metal sulfide.
12. The process of Claim 11 wherein at least one subsequent flotation step produces a lower grade desired metal concentrate.
13. The process of Claim 1 wherein at least a portion of said selected metal ion rich aqueous stream is produced by solubilizing desired metal ions from a quantity of said mineral ore concentrate in a dissolution zone to produce at least a portion of said desired metal ion rich aqueous stream.
14. The process of Claim 13 wherein said mineral ore concentrate charged to said process is mechanically split into a first portion which is passed to said dissolution zone and a second portion which is passed to said reaction zone.
15. The process of Claim 13 wherein said selected metal ions are solubilized in an acidic stream containing sulfuric acid or sulfate compounds at a pH from about 0.8 to about 1.2.
16. The process of Claim 1 wherein said mineral ore concentrate charged to said process is selectively separated by a froth flotation process into a desired metal rich portion and a desired metal lean portion, wherein said desired metal rich portion is passed to a dissolution zone where desired metal is solubilized from said desired metal rich portion to produce at least a portion of said desired metal ion rich aqueous solution and wherein said desired metal lean portion is passed to said reaction zone.
17. The process of Claim 16 wherein said desired metal is copper and wherein iron is present in said concentrate.
18. The process of Claim 16 wherein said iron is present in an amount greater than said copper.
19. The process of Claim 16 wherein said desired metal ions are solubilized in an acidic stream containing sulfuric acid or sulfate compounds at a pH from about 0.8 to about 1.2.
20. The process of Claim 1 wherein at least a portion of said desired metal ion rich aqueous stream is produced by solubilizing desired metal ions from a quantity of said mineral ore concentrate into an aqueous solution containing ammonia at a pH from about 9.5 to about 10.5.
21. The process of Claim 20 wherein said mineral ore concentrate contains quantities of chalcocite coated pyrite.
22. The process of Claim 1 wherein said reaction zone comprises a plurality of reaction zones in series.
23. The process of Claim 22 wherein a portion of said desired metal contained in said mineral ore concentrate is solubilized by oxidizing a portion of said desired metal in said mineral ore concentrate in at least one first reaction zone to produce at least a portion of said metal ion rich aqueous stream in situ in said reaction zone.
24. The process of Claim 23 wherein said reaction to produce said product aqueous slurry is accomplished in subsequent reaction zones.
25. The process of Claim 23 wherein oxygen is charged to said first reaction zone in an amount sufficient to oxidize from about 30 to about 70 percent of the desired metal initially present in said mineral ore concentrate.
26. The process of Claim 1 wherein said desired metal ion rich stream is produced by solubilizing desired metal from a process recycle stream.
27. A process for selectively enriching a sulfide mineral ore concentrate in a desired metal, said sulfide mineral ore concentrate being contained in an aqueous slurry of finely divided particulate solids containing desired metal sulfides and mixed metal sulfides containing said desired metal and other metal or other metal sulfides, said other metal having a lower affinity for sulfur than said desired metal to produce an enriched concentrate having an increased desired metal content; said process comprising:
(a) separating said sulfide mineral ore concentrate into a first portion and a second portion;
(b) directing said first portion of said sulfide mineral ore concentrate to a dissolution zone wherein desired metal is solubilized from said first portion of said sulfide mineral ore concentrate to produce a desired metal ion rich aqueous stream;
(c) reacting said desired metal ion rich aqueous stream with said second portion of said sulfide mineral ore concentrate in a reaction zone to produce a product aqueous slurry containing increased quantities of desired metal sulfides;
(d) separating said product aqueous slurry into a desired metal ion depleted aqueous stream and a finely divided particulate solids stream containing said desired metal sulfides; and, (e) separating at least a portion of said desired metal sulfides from said finely divided particulate solids stream by a froth flotation process to produce said enriched concentrate.
(a) separating said sulfide mineral ore concentrate into a first portion and a second portion;
(b) directing said first portion of said sulfide mineral ore concentrate to a dissolution zone wherein desired metal is solubilized from said first portion of said sulfide mineral ore concentrate to produce a desired metal ion rich aqueous stream;
(c) reacting said desired metal ion rich aqueous stream with said second portion of said sulfide mineral ore concentrate in a reaction zone to produce a product aqueous slurry containing increased quantities of desired metal sulfides;
(d) separating said product aqueous slurry into a desired metal ion depleted aqueous stream and a finely divided particulate solids stream containing said desired metal sulfides; and, (e) separating at least a portion of said desired metal sulfides from said finely divided particulate solids stream by a froth flotation process to produce said enriched concentrate.
28. The process of Claim 27 wherein at least one process recycle stream is directed to said dissolution zone.
29. The process of Claim 27 wherein said desired metal is selected from the group consisting of copper, nickel, silver, gold, lead and zinc.
30. The process of Claim 27 wherein said desired metal is copper.
31. A process for selectively enriching a sulfide mineral ore concentrate in a desired metal, said sulfide mineral ore concentrate being contained in an aqueous slurry of finely divided particulate solids containing desired metal sulfides and mixed metal sulfides containing said desired metal and other metal or other metal sulfides, said other metal having a lower affinity for sulfur than said desired metal to produce an enriched concentrate having an increased desired metal content; said process comprising:
(a) selectively separating said sulfide mineral ore concentrate into a desired metal rich portion and a desired metal lean portion;
(b) directing said desired metal rich portion to a dissolution zone wherein desired metal is solubilized from said desired metal rich portion to produce a desired metal ion rich aqueous stream;
(c) reacting said desired metal ion rich aqueous stream with said desired metal lean portion of said sulfide mineral ore concentrate in a reaction zone to produce a product aqueous slurry containing increased quantities of desired metal sulfides;
(d) separating said product aqueous slurry into a desired metal ion depleted aqueous stream and a finely divided particulate solids stream containing said desired metal sulfides; and, (e) separating at least a portion of said desired metal sulfides from said finely divided particulate solids stream by a froth flotation process to produce said enriched concentrate.
(a) selectively separating said sulfide mineral ore concentrate into a desired metal rich portion and a desired metal lean portion;
(b) directing said desired metal rich portion to a dissolution zone wherein desired metal is solubilized from said desired metal rich portion to produce a desired metal ion rich aqueous stream;
(c) reacting said desired metal ion rich aqueous stream with said desired metal lean portion of said sulfide mineral ore concentrate in a reaction zone to produce a product aqueous slurry containing increased quantities of desired metal sulfides;
(d) separating said product aqueous slurry into a desired metal ion depleted aqueous stream and a finely divided particulate solids stream containing said desired metal sulfides; and, (e) separating at least a portion of said desired metal sulfides from said finely divided particulate solids stream by a froth flotation process to produce said enriched concentrate.
32. The process of Claim 31 wherein said desired metal is selected from the group consisting of copper, nickel, silver, gold, lead and zinc.
33. The process of Claim 32 said mineral ore concentrate is a copper-iron-sulfide concentrate.
34. A process for selectively enriching a sulfide mineral ore concentrate in a desired metal said sulfide mineral ore concentrate being contained in an aqueous slurry of finely divided particulate solids containing desired metal sulfides and mixed metal sulfides containing a desired metal and other metal or other metal sulfides, said other metal having a lower affinity for sulfur than said desired metal to produce an enriched concentrate having an increased desired metal content; said process comprising:
(a) directing said sulfide mineral ore concentrate to a multistage reaction zone;
(b) solubilizing a portion of said desired metal in said sulfide mineral ore concentrate in a first stage of said multistage reaction zone to produce a desired metal ion rich aqueous solution;
(c) reacting said desired metal ion rich aqueous stream with said sulfide mineral ore concentrate in a subsequent stage of said multistage reaction zone to produce a product aqueous slurry containing increased quantities of desired metal sulfides;
(d) separating said product aqueous slurry into a desired metal ion depleted aqueous stream and a finely divided particulate solids stream containing said desired metal sulfides; and, (e) separating at least a portion of said desired metal sulfides from said finely divided particulate solids stream by a froth flotation process to produce said enriched concentrate.
(a) directing said sulfide mineral ore concentrate to a multistage reaction zone;
(b) solubilizing a portion of said desired metal in said sulfide mineral ore concentrate in a first stage of said multistage reaction zone to produce a desired metal ion rich aqueous solution;
(c) reacting said desired metal ion rich aqueous stream with said sulfide mineral ore concentrate in a subsequent stage of said multistage reaction zone to produce a product aqueous slurry containing increased quantities of desired metal sulfides;
(d) separating said product aqueous slurry into a desired metal ion depleted aqueous stream and a finely divided particulate solids stream containing said desired metal sulfides; and, (e) separating at least a portion of said desired metal sulfides from said finely divided particulate solids stream by a froth flotation process to produce said enriched concentrate.
35. The process of Claim 34 wherein said desired metal is selected from the group consisting of copper, nickel, silver, gold, lead and zinc.
Applications Claiming Priority (2)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| US67635584A | 1984-11-29 | 1984-11-29 | |
| US676,355 | 1984-11-29 |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CA1258181A true CA1258181A (en) | 1989-08-08 |
Family
ID=24714186
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CA000477505A Expired CA1258181A (en) | 1984-11-29 | 1985-03-26 | Process for producing enriched mineral ore concentrates |
Country Status (1)
| Country | Link |
|---|---|
| CA (1) | CA1258181A (en) |
Cited By (3)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US9587290B2 (en) | 2013-03-14 | 2017-03-07 | Orway Mineral Consultants (Wa) Pty, Ltd. | Hydrometallurgical method for the removal of radionuclides from radioactive copper concentrates |
| US10407753B2 (en) * | 2014-12-19 | 2019-09-10 | Flsmidth A/S | Methods for rapidly leaching chalcopyrite |
| CN112619878A (en) * | 2020-11-10 | 2021-04-09 | 西北矿冶研究院 | Comprehensive recovery process for iron symbiotic nonferrous metal copper, lead and zinc |
-
1985
- 1985-03-26 CA CA000477505A patent/CA1258181A/en not_active Expired
Cited By (4)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US9587290B2 (en) | 2013-03-14 | 2017-03-07 | Orway Mineral Consultants (Wa) Pty, Ltd. | Hydrometallurgical method for the removal of radionuclides from radioactive copper concentrates |
| US10407753B2 (en) * | 2014-12-19 | 2019-09-10 | Flsmidth A/S | Methods for rapidly leaching chalcopyrite |
| CN112619878A (en) * | 2020-11-10 | 2021-04-09 | 西北矿冶研究院 | Comprehensive recovery process for iron symbiotic nonferrous metal copper, lead and zinc |
| CN112619878B (en) * | 2020-11-10 | 2023-01-03 | 西北矿冶研究院 | Comprehensive recovery process for iron symbiotic nonferrous metal copper, lead and zinc |
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