US3728430A - Method for processing copper values - Google Patents

Method for processing copper values Download PDF

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US3728430A
US3728430A US00097851A US3728430DA US3728430A US 3728430 A US3728430 A US 3728430A US 00097851 A US00097851 A US 00097851A US 3728430D A US3728430D A US 3728430DA US 3728430 A US3728430 A US 3728430A
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copper
ore
percent
leaching
sulfur
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J Clitheroe
F Lacy
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ANLIN CO
ANLIN CO OF NEW JERSEY US
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    • CCHEMISTRY; METALLURGY
    • C01INORGANIC CHEMISTRY
    • C01GCOMPOUNDS CONTAINING METALS NOT COVERED BY SUBCLASSES C01D OR C01F
    • C01G3/00Compounds of copper
    • C01G3/12Sulfides
    • BPERFORMING OPERATIONS; TRANSPORTING
    • B03SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03BSEPARATING SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS
    • B03B1/00Conditioning for facilitating separation by altering physical properties of the matter to be treated
    • B03B1/04Conditioning for facilitating separation by altering physical properties of the matter to be treated by additives

Abstract

The steps of leaching copper values from oxide ores, mixed sulfide-oxide ores, and other oxide copper bearing materials in an aqueous medium or suspension containing soluble sulfites or bisulfites and precipitating the copper as copper sulfide in an aqueous medium or suspension containing elemental sulfur and soluble sulfites or bisulfites. The precipitated copper, along with any other copper present as elemental copper or copper sulfide, is then separated by conventional methods.

Description

Unite States Patent 1191 Clitheroe et a1.
[ 1 Apr. 17, 1973 METHOD FOR PROCESSING COPPER VALUES [75] Inventors: Jay B. Clitheroe, Salt Lake City, Utah; Forrest H. Lacy, Jr., Houston,
21 Appl. No.: 97,851
[52] US. Cl. ..423/26, 423/27, 423/37, 423/43, 75/101 R, 75/108, 75/2, 299/5 [51] Int. Cl ..C0lg 3/12, C22b 15/08 [58] Field of Search ..75/117, 101 R, 108, 75/2; 23/135; 299/5; 423/26, 27, 37, 43'
[56] References Cited UNITED STATES PATENTS 1,335,001 3/1920 Hovland et a1. ..75/2 1,197,589 9/1916 Bacon ..75/2 2,332,145 10/1943 Hay ..23/.l35 3,322,532 5/1967 Wieder.... ..75/l17 X 3/1920 Hovland et a1. ..75/2
3,168,396 2/1965 Barker ..75/1 17 X 3,573,896 4/1971 Hori ..75/117 3,218,161 11/1965 Kunda' et a1. .....75/1l7 X 3,330,649 7/1967 Welsh ...75/2 X 1,679,294 7/1928 Dietzsch. .....'75/1 17 1,360,666 1 1/1920 Mills ..75/117 X Primary ExaminerG: T. Ozaki Attorney-Paul E. Harris, Lee R. Larkin and Marcus L. Thompson ABSTRACT The steps of leaching copper values from oxide ores, I
4 Claims, 1 Drawing Figure METHOD FOR PROCESSING COPPER VALUES BACKGROUND OF THE INVENTION 1. Field of the Invention The field of the invention is methods for hydrometallurgical-physical beneficiation of metalliferous ores, mixed sulfide-oxide ores, and middlings, slags, tailings, and dumps containing oxide copper values.
2. Description of the Prior Art Ores of sufficiently high copper content may be smelted directly, but lower grade copper ores, dumps, middlings, and slag usually must be concentrated before shipment or smelting in order to economically obtain the copper values therefrom.
Ores high in copper sulfide content quite often contain a portion of oxide ores intermingled in the naturally occurring deposits. The term oxide is meant to describe the nonsulfide ore constituents, e.g., copper oxides, silicates, carbonates, hydroxides, and sulfates. Rich sulfide ores are commonly concentrated by grinding and flotation processes which, in many cases, allow copper oxides to pass substantially unchanged through the process to waste.
Certain oxide-containing ores are presently being concentrated by leaching with sulfuric acid and precipitating copper or copper sulfide by the use of ground or powdered iron or shredded scrap or hydrogen sulfide. In many cases these steps are followed with separation of the precipitate of cement copper or copper sulfide by conventional froth flotation. Such a technique is familiarly known as LPF. Conventional LPF has many disadvantages, including the following:
a. The reagents used in conventional LPF are relatively expensive.
b. Some of the iron is consumed by the required free acid and by oxidation and thus wasted.
c. Some of the soluble copper tends to plate out on coarser particles of iron, thus wasting iron and sometimes rendering the coarse particles of plated copper non-floatable.
d. Copper sulfide precipitated by hydrogen sulfide tends to be colloidal in nature and therefore difficult to concentrate by flotation, especially in the presence of slimes. In addition, the use of an extremely toxic substance such as hydrogen sulfide poses many safety engineering problems.
e. The sulfuric acid leach is often slow.
f. Acid consuming gangue often requires excessive quantities of relatively expensive sulfuric acid.
Even though conventional LPF has been widely used, copper oxide wastes and many copper oxide ores, some in relatively large deposits, have been difficult or impossible to process economically. This is especially true of .those ores containing a large proportion of slimes, copper silicates or acid-consuming constituents.
BRIEF DESCRIPTION OF THE DRAWING The drawing is a flow sheet showing one embodiment of the invention preferred by the inventor for processing materials containing oxide copper.
SUMMARY OF THE INVENTION This invention improves oxide copper beneficiation processes and includes the steps of leaching the oxide copper values from metalliferous materials in the presence of water and soluble sulfites or bisulfites and precipitating the copper values as copper sulfide in the presence of elemental sulfur and soluble sulfites or soluble bisulfites. The method may include the additional step of separating the dissolved copper from the undissolved o're prior to the precipitation 'of copper sulfide. After precipitation, the copper sulfide is then separated from the aqueous medium.
In the preferred embodiment of the invention, the leaching and precipitating steps are carried out substantially simultaneously in a pulp containing soluble sulfites or soluble bisulfites and elemental sulfur. The precipitate is then separated from the pulp, preferably by conventional froth flotation.
Another. embodiment of the invention includes the steps of leaching the oxide copper containing material in situ with a solution containing soluble sulfites or bisulfites, transferring the leached solution containing copper values to a precipitation vessel, and precipitating the copper values as copper sulfide in the presence of soluble sulfites and bisulfites and elemental sulfur.
Certain embodiments of the invention include the step of leaching the oxide copper in the presence of water, soluble chloride, and a material selected from the group consisting of soluble bisulfites and sulfur dioxide. I
DESCRIPTION OF THE PREFERRED EMBODIMENT The chemistry of the embodiment preferred by the inventor may be schematically represented as follows:
This-embodiment of the invention is practiced by slurrying the copper-containing material with water and reducing the material to a suitable degree of fineness before entering the leach-precipitation circuit. However, any method of exposing the copper values of the raw material to the leaching-precipitation medium is acceptable. The slurry thus formed is referred to as pulp. Elemental sulfur and water may be added prior to or after the grinding operation, and if added prior to grinding, the sulfur would, of course, be ground along with the raw material. The water need not be fresh water. It may be salt water and, under certain circumstances, more desirably may be salt water. Salt water is defined as water which naturally contains sodium chloride or water to which sodium chloride has been added. Such elemental sulfur maybe introduced into the leach-precipitation circuit at any point prior to the precipitation step. From the grnding operation the pulp flows to leach-precipitation agitators. Soluble sulfite furnished in the form of hot gas containing sulfur dioxide is contacted with the pulp, both heating the mixture and furnishing sulfur dioxide for sulfurous acid. In the drawing, a series of leach-precipitation agitators are shown for contacting the pulp and hot S0 bearing gas. Soluble sulfates may be added to the leachprecipitation agiator as accelerators. Oxide copper is leached (as shown in equation [1]) by the sulfurous acid formed by hydrolysis of the sulfur dioxide bubbled through the slurry (as shown in equation [2]). A portion of the elemental sulfur in the pulp serves as the precipitant, yielding a precipitate of copper sulfide (as shown in equation [3]). It has been found beneficial to include in the process an excess of elemental sulfur above that amount which stoichiometrically reacts with the leached copper. Some of the copper sulfide clings to excess free sulfur particles present during the precipitation step. As may be seen from equation (3),
sulfuric acid is produced in the leach-precipitation reaction and is available to react with all or a part of any acid-consuming constituents in the ore, thereby lowering the chemical requirements if the gangue is acid consuming.
It is believed that a preponderance of the copper, represented by the Cu symbol in equations (l)(3), is present in the precipitate as the cupric form. However, there 'maybe some copper present as cuprous sul-' fide after precipitation, depending on the specific ore being processed. It is of no consequence to the operation of the invention which valence state the copper takes. r
The elemental sulfur should be ground to, or supplied in, a fine particle size, because its reaction rate is somewhat dependent upon its available surface area. The preferred method for introducing the sulfur is to grind it directly along with the ore being processed. The source of sulfur may be in a number of forms, e.g., solid elemental sulfur, molten sulfur sprayed onto the ore, or even naturally-occurring sulfur ores, many of which such sulfur ores are not otherwise economically processable. The elemental sulfur for the precipitation could be supplied by a reduction roast of pyrite.
The precipitate is then physically separated, and in this preferred embodiment, is floated by conventional techniques. Many of the particles of copper sulfide surround and adhere mom or more particles of elemental sulfur and thus are easily floated and cleaned. The precipitate of copper sulfide was found to be coarse and surprisingly easy to float from the pulp. No colloidal precipitate, as found in H 8 precipitations, or adherence to iron particles, as in iron precipitations, was
noted. The ease of recovery. by flotation contributes to both the economy of the process and its suitability for marginal ores.
Other materials containing copper are also floated with the copper sulfide precipitate in this embodiment. Most of the excess sulfur not consumed in the process is recovered in the flotation concentrate, the flotation being carried out after the pulp-to-pulp heat exchange step shown in the drawing. Naturally occurring copper sulfide, as well as native copper, if present, are floated with the precipitate and recovered in the froth flotation step. It has been discovered that'the process brightens (detarnishes) copper sulfide and native copper particles originally present in the ore, thereby improving their flotation characteristics. Some native copper, if present, may be leached and'precipitated in theleachprecipitation agitator. This leached and precipitated native copper will appear along with leached and precipitated oxide copper as copper sulfide.
The concentrate, consisting primarily of element sulfur and copper sulfide, is dewatered and roasted to yield heat and sulfur dioxide for circulation to the leach-precipitation agitators. The roaster may be operated so that only the elemental sulfur 'in the concentrate is oxidized or it may be operated so that the copper sulfide in the concentrate is also oxidized, yielding an excellent driedconcentrate which maybe transferred to a smelter or refinery.
The leach-precipitation agitators may be provided with a vent to discharge any excess gaseous sulfur dioxide at an altitude which will comply with any air pollution regulations, but preferably are provided with a conduit to transfer any excess gaseous sulfur dioxide to a basic metallic earth contact stage. Such a contactor would contain a basic metallic earth material such as sodium or potassium oxide, hydroxide, or carbonate, calcium or magnesium oxide, hydroxide, or carbonate, ammonium hydroxide or carbonate, or ferrous hydroxide, oxide, or carbonate. Such a contactor could also contain sodium or potassium chloride, calcium or magnesium chloride, ammonium chloride, and ferrous or ferric chloride. Naturally occurring'contactor materials could be dolomite, calcite, soda ash, or limestone or ordinary salt. There should be present in any such contact stage some water, preferably that which would later be used in the grinding and the leach-precipitation steps, all or a portion of which may come from the tailings pond. Such a contact stage can provide a source of soluble fixed sulfites and bisulfites or chloride to the process, all of which have been demonstrated to be beneficial. In addition, utilization of the waste gases in this manner is beneficial not only as a source of soluble fixed sulfites and bisulfites and chlorides, but if the materials in the contactor are properly proportioned, this virtually eliminates the'problem of venting any sulfur dioxide to the atmosphere, a pollution problem which has plagued various metal processing installations in the past. Soluble fixed sulfites and bisulfites may be defined as any source of soluble sulfite or bisulfite ions other than sulfurous acid. Sulfurous acid would be included in the group defined as soluble sulfites and soluble bisulfites, which group also includes the sulfites and bisulfites of sodium, calcium, ammonium, magnesium, and potassium, as well as the sulfite of iron. Solublechlorides are defined as any water soluble chloride-containing material which ionizes to furnish the chloride in water solution.
Some or all of the heating and all or a part of the sulfur dioxide for leaching could be supplied by utilizing hot smelter'gases containing sulfur dioxide or by other sulfur dioxide bearing stack or vent gases. Heat and sulfur dioxide also could be supplied simply by burning elemental sulfur with or without the roasting of sulfurbearing flotation concentrates. Soluble sulfite or bisulfite, in the form of sulfur dioxide, for leaching could be supplied by an oxidation roast of the FeS resulting from a reduction roast of pyrite or could,of course, be supplied as a solution of sulfurous acid.
It should be noted that in the preferred embodiment sulfur dioxide introduced into the process from a roaster gas, smelter gas or other sulfur dioxide bearing vent or stack gas does not have to be in a clean gas stream. Such gas may contain substantial amounts of normally objectionable particulate matter '(including copper or its compounds) and may even contain normally harmful or objectionable elements, such as tellurium, selenium, etc. In the preferred embodiment such materials end up either as a part of the recovered concentrate, in the case of most of the copper, or as harmlessly fixed materials in the tailings. Thus, both air and water pollution is minimized or eliminated.
The process may be carried out at a temperature ranging from about 80 to 212 F. at atmospheric pressure and at a pH of 1.0 to 9.4. However, a temperature of about 130 F to about 170 F. and a pH of about 1.0 to about 6.0 has been demonstrated to be practical, economical and desirable.
With the addition of pressure equipment at the leach-precipitation stage, the process may be carried out above boiling. A convenient source of heat to supply such elevated heat requirements would make higher temperature extraction and precipitation with pressure equipment attractive as an alternative embodiment of the process.
A part of the heat requirements may be efficiently achieved by indirect heat exchange of the final leach pulp with the pulp entering the leach contactor. The sulfur dioxide and all or a part of the heat requirements are obtained most easily by burning sulfur directly or roasting the copper sulfide concentrate which will nor.- mally contain some free sulfur. In some cases, roasting the concentrate will not provide enough heat or enough sulfur dioxide to reach optimum reaction conditions. In that case, additional sulfur, natural gas, or fuel oil may be burned to supply additional heat, and, if sulfur is burned, additional sulfur dioxide. The amount of sulfur dioxide and heat supplied by the hot gas stream bubbled into the leach-precipitation agitators may vary with the ore being processed.
In the case of an ore containing an oxide mineral that is somewhat refractory with respect to sulfurous acid, the low-cost sulfurous acid could be used in a primary leach-precipitation step to digest and precipitate the easily dissolved metal minerals, while being assisted to leach and precipitate such refractory minerals as a secondary leaching step by the addition of an economically small quantity of more valuable mineral acid, such as sulfuric, hydrochloric, hydroflouric, nitric, hydroiodic, fluosilicic, phosphoric, acetic and formic. Additionally, it has been demonstrated that the addition of a soluble chloride in the presence of sulfur dioxide or a soluble bisulfite aids in the leaching portion of the process. The amount of such acid or soluble chloride added or used will vary with the particular ore being processed, and the acid, if added, should be present in an amount sufficient to keep the leachprecipitation reaction within optimum parameters.
The following examples illustrate the process as v hereinbefore disclosed:
EXAMPLE 1 Red Bed copper ore from North Texas, containing about 85 percent illite clays and assaying approximately 1.1 percent copper, of which 67 percent was in oxide" form, and containing the equivalent of 1.5 percent calcium carbonate, was put into a ball mill with the equivalent of 15 pounds per ton of elemental sulfur and wet ground with water to approximately minus 200 mesh at about a 45 percent solids level. All percentages and ratios are calculated on weight basis. The finely ground elemental sulfur and copper ore were mixed with the equivalent of pounds sulfur dioxide per ton of ore in the form of sulfurous acid and agitated for .20 minutes at a pH of 2.6-2.9 at 138 F. Conventional froth flotation was then carried out. Copper sulfide, some of which adhered to particles of elemental sulfur, was then recovered by conventional froth flotation. The cleaner concentrate contained 76.95 percent of the copper originally in the ore with 7.51 percent of the copper remaining-in the cleaner tail and 15.54 percent of the copper remaining in the rougher tail.
Both heat and low pH contribute to shorter leach- As a comparison, a different sample of the same ore was put into a ball mill and ground to minus 100 mesh with water at a percent solids level of approximately 50 percent. It was then processed by conventional LPF techniques, e.g., leached for 40 minutes with the equivalent of pounds of sulfuric acid per ton of ore at a pH of 1.6, precipitated for 20 minutes with the equivalent of 80 pounds of powdered iron per ton of ore (of which 27.3 pounds of iron was consumed) and then treated for 20 minutes by conventional froth flotation with two stages of recleaning. The result of this test showed the distribution of the copper to be 65.4 percent in the LPF concentrate, 15.8 percent in the cleaner tail, 5.2 percent'in the recleaner tailing, and 13.6 percent remaining in the rougher tail. This test represented the best results obtained from an extensivetesting program of conventional LPF as applied to this ore.
EXAMPLE 2 The Red Bed copper ore of Example 1 was wet ground with water and elemental sulfur in an amount equivalent to 30 pounds per ton of ore and digested for 20 minutes at a temperature of 146 F. with the addition of the equivalent of ten pounds of ammonium sulfate per ton of ore in the leach-precipitation agitator. During this test the equivalent of 80 pounds per ton of sulfur dioxide was added to the leach-precipitation agitator. The pH of the leach-precipitation slurry remained at 2.7 during leaching-precipitating. The cleaner concentrate showed a recovery of 88.31 percent of the total copper originally in the ore with 2.89 percent in the cleaner tail and 8.80 percent in the rougher tail.
EXAMPLE 3 The Red Bed copper ore of Examples 1 and 2 with the equivalent of 30 pounds elemental sulfur per ton of ore was ground to about a minus 200 mesh with water at a percent solids level of approximately 33 percent.
EXAMPLE 4 The Red Bed copper ore or examples 1 through 3, treated with the equivalent of 30 pounds elemental sulfur per tom of ore, was ground to about minus 200 mesh with waterat a percent solids level of approximately 33 percent. The resulting slurry was mixed with the equivalent of 30 pounds of sulfur dioxide per ton of ore and digested for 20 minutes at 80 F., the approxi- Other examples demonstrating the efficacy of this embodiment of the invention have been performed over a pH range of about 1.0 to about 9.4, at temperatures from 80 F. to about 2 1 2 F .,'and at elemental sulfur to copper molar ratios from 0.5 to 4.0. Sulfur dioxidc to copper molar ratios between 0.5 and 4.5, and soluble fixed sulfite and bisulfite to copper molar ratios between 0.5 and 4.5 have been used in other examples and shown to be operative.
EXAMPLE 7 Uinta Mountain sandstone ore from Utah containing azurite and malachite with a copper content of 2.85 percent and a siliceous matrix was wet ground with the equivalent of 80 pounds sulfur per ton of ore and water v to about minus 65 mesh at a pulp density of 39.2 percent solids. Eighty pounds of sulfur dioxide per ton of ore and 50 pounds per ton of soluble fixed sulfite salts were digested with the slurry at 170 F. for 20 minutes at a pH of 1.5 to 1.7. The concentrate obtained by conmate ambient temperature. The leaching-precipitation slurry contained the equivalent of 50 pounds of soluble sulfite salts per ton of ore. The beginning pH of the leach-precipitation slurry was 3.1, and the ending pH was 4.2. The recovery by conventional flotation was 67.32 percent of the copper contained in the-original ore sample with 2.33 percent of the copper remaining in the cleaner tail and 30.35 percent remaining in the rougher tail.
EXAMPLE 5 The Red Bed copper ore of examples 1 through 4 was ground with the equivalent of pounds of elemental sulfur per ton of ore to about minus 200 mesh with water at a percent solids level of approximately 33 percent. The resulting slurry was mixed with the equivalent of 110 pounds of soluble fixed sulfite salts per ton of ore and digested for 20 minutes at 160 F. The beginning pH of the slurry was 8.4and the ending pH was 9.4. The recovery by conventional flotation was 45.03 percent of the copper contained in the original ore sample with 1.94 percent of the copper remaining in the cleaner tail and 53.03 percent remaining in the rougher tail.
EXAMPLE 6 the equivalent of 73 pounds of sodium chloride per ton of ore. The recovery by conventional flotation was 96.20 percent of the copper contained in the original ore sample with 0.58 percent of the copper remaining in the cleaner tail and 3.22 percent remaining in the rougher tail.
ventional froth flotation in this example contained 93.05 percent of the total copper in the ore sample with the cleaner tail containing 0.54 percent and a rougher tail containing 6.41 percent of the total copper.
EXAMPLE 8 A conglomerate oxide ore from the Philippines, with a 2.59 percent copper content and containing chrysocolla and tenorite in a siliceous matrix, was
ground with the equivalent of 104 pounds of sulfur'per ton of ore and water to a minus 65 mesh. The equivalent of l0 pounds of ammonium sulfate per ton of ore and 77 pounds of sulfur dioxide per ton of ore were digested for 60 minutes at a temperature of 160 F. at a pH of about 1.6. The leach-precipitation slurry additionally contained the equivalent of 100 pounds of soluble fixed sulfite salts per ton of ore. In this example,
a seconary leach-precipitation step was conducted,.
consisting of adding the equivalent of 50 pounds of sulfuric acid per ton of ore to the leach-precipitator and digesting-precipitating for an additional 20 minutes at a temperature of 160 F. and a pH of 1.1 to 1.6. The concentrate obtained using conventional froth flotation contained 91.06 percent of the copper in the original ore, with a cleaner tail containing 0.54 percent and a rougher tail containing 7.95 percent of the original copper.
A comparison test for leaching only as performed in the prior art, showed that this ore ground to the same fineness yielded only 85 percent of its total copper in a 5 percent sulfuric acid solution after four hours of conventional leaching.
EXAMPLE 9 An Arizona Strip oxide ore containing quartz, approximately 20 percent calcium carbonate, azurite, and malachite, for conventional leaching, required 500 pounds of H 80 per ton of ore yielding percent copper extraction in the leaching stage only, without fide, would be experienced with conventional recovery techniques. The same Arizona ore was treated in accordance with the preferred method by grinding with the equivalent of 117 pounds of elemental sulfur per ton of ore and water to minus 65 mesh. The resulting presently preferred slurry was then treated with the equivalent of 348 pounds of sulfur dioxide per ton of ore and the equivalent of 50 pounds of soluble fixed sulfite salts per ton of ore for 40 minutes at 160 F. at a'pH of 3.8. The concentrate obtained using conventional froth flotation contained 86.29 percent of the copper originally in the ore, with a cleaner tail containing 0.56 percent and a rougher tail containing 13.12 percent of the original copper.
As an illustration of processing cost savings affected by the preferred embodiment, reagent costs for processing the ore of this example are less than half of the cost of leaching only by the methods of the prior art.
The concentration of various reagents used in the process will vary, of course, depending on the specific ore being processed. Maximum efficiency of the process may be achieved by varying the time, temperature and pH of the leaching, precipitating and separating steps in accordance with generally known scientific principles. For example, a high-silica or high-lime ore may require the lowering of the pH of the leachingprecipitating medium or increasing the reaction temperature for optimum results.
In addition to adaptation of the described embodiments for normal ore processing techniques, one em bodiment of the invention includes working over subterranean deposits by leaching ores in situ. Heat and pressure present at such underground mines might be utilized containing the leaching process. The pregnant skilled in the art the manner of carrying out the invention. it is to be understood that the form of the invention herewith shown and described is to be taken as the embodiment. For example, equivalent sequences of steps and process configurations may be substituted for those illustrated and described herein, and certain features of the invention may be utilized independently of the use of other features, all as would be apparent to one skilled in the art after having the benefit of this description of the invention.
What is claimed is:
1. In a method for processing oxide ores, mixed sultide-oxide ores and middlings, dumps and slag contain ing oxide copper values, the combination of steps comprising:
leaching and substantially simultaneously precipitating said oxide copper values in said material at a pH of about 1.0 to about 9.4 and at a temperature of from about F to about 212 F in an a ueous pulp with elemental sulfur and a material se ected from the group consisting of soluble sulfites and solublebisulfites, whereby said copper is precipitated as copper sulfide; and,
separating copper sulfide from said pulp by froth flotation.
2. The method of claim 1 wherein:
said leaching and precipitating step is carried out in the presence of a stoichiometric excess of sulfur dioxide and elemental sulfur in relation to the amount of oxide copper in said copper containing material.
3. The method of claim 1 wherein:
said leaching and precipitating step includes the additional step of adding an acid selected from the group consisting of sulfuric, hydrochloric, nitric, hydroiodic, hydrofluoric, fluosilicic, phosphoric, acetic and formic to said pulp prior to completion of the precipitating portion of said leaching and precipitating step in an amount sufficient to maintain the pH thereof in the range of from about 1.0 to about 9.4.
4. The method of claim 1 wherein:
said leaching and precipitating step is carried out in said pulp in the presence of a soluble chloride.

Claims (3)

  1. 2. The method of claim 1 wherein: said leaching and precipitating step is carried out in the presence of a Stoichiometric excess of sulfur dioxide and elemental sulfur in relation to the amount of oxide copper in said copper containing material.
  2. 3. The method of claim 1 wherein: said leaching and precipitating step includes the additional step of adding an acid selected from the group consisting of sulfuric, hydrochloric, nitric, hydroiodic, hydrofluoric, fluosilicic, phosphoric, acetic and formic to said pulp prior to completion of the precipitating portion of said leaching and precipitating step in an amount sufficient to maintain the pH thereof in the range of from about 1.0 to about 9.4.
  3. 4. The method of claim 1 wherein: said leaching and precipitating step is carried out in said pulp in the presence of a soluble chloride.
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Cited By (15)

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US3834760A (en) * 1973-07-18 1974-09-10 Kennecott Copper Corp In-situ generation of acid for in-situ leaching of copper
US3894770A (en) * 1974-06-12 1975-07-15 Kennecott Copper Corp Wellbore oxidation of lixiviants
US4006014A (en) * 1975-07-28 1977-02-01 Canadian Industries Limited Use of tetraalkylammonium halides as flotation collectors
US4066520A (en) * 1976-09-01 1978-01-03 Envirotech Corporation Slurry electrowinning process
US4091070A (en) * 1976-08-25 1978-05-23 Inspiration Consolidated Copper Company Recovery of copper
FR2602797A1 (en) * 1986-07-29 1988-02-19 Khim Metall Institu PROCESS FOR ENRICHING DIFFICULTLY ENRICHABLE OXIDE COPPER ORE
US4735783A (en) * 1987-04-22 1988-04-05 Falconbridge Limited Process for increasing the selectivity of mineral flotation
DE3690783C2 (en) * 1986-10-04 1990-11-22 Shimiko Metallurg I Central Nc Oxide copper mineral concn. by flotation - involving pre-sulphidation with molten sulphur
US5178842A (en) * 1986-03-10 1993-01-12 Outokumpu Oy Method for precipitating and separating metals
US5443622A (en) * 1994-02-28 1995-08-22 Kennecott Corporation Hydrometallurgical processing of impurity streams generated during the pyrometallurgy of copper
WO1996025361A1 (en) * 1995-02-17 1996-08-22 Baker Hughes Incorporated Copper precipitation process
US5616168A (en) * 1994-02-28 1997-04-01 Kennecott Utah Copper Corporation Hydrometallurgical processing of impurity streams generated during the pyrometallurgy of copper
US20090241734A1 (en) * 2008-03-25 2009-10-01 Imagawa Harue Method of leaching copper sulfide ores containing chalcopyrite
WO2015081385A1 (en) * 2013-12-03 2015-06-11 The University Of Queensland Copper processing method
US20170191144A1 (en) * 2014-05-28 2017-07-06 Matthew Leslie Sutcliffe Method for ammoniacal leaching of copper from oxidised copper ores

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US3834760A (en) * 1973-07-18 1974-09-10 Kennecott Copper Corp In-situ generation of acid for in-situ leaching of copper
US3894770A (en) * 1974-06-12 1975-07-15 Kennecott Copper Corp Wellbore oxidation of lixiviants
US4006014A (en) * 1975-07-28 1977-02-01 Canadian Industries Limited Use of tetraalkylammonium halides as flotation collectors
US4091070A (en) * 1976-08-25 1978-05-23 Inspiration Consolidated Copper Company Recovery of copper
US4066520A (en) * 1976-09-01 1978-01-03 Envirotech Corporation Slurry electrowinning process
US4096053A (en) * 1976-09-01 1978-06-20 Envirotech Corporation Slurry electrowinning apparatus
US5178842A (en) * 1986-03-10 1993-01-12 Outokumpu Oy Method for precipitating and separating metals
AU604006B2 (en) * 1986-07-29 1990-12-06 Gosudarstvenny Institut Po Proektirovaniju Predpriyaty Tsvetnoi Metallurgii Method of concentration of refractory oxidized copper ores
FR2602797A1 (en) * 1986-07-29 1988-02-19 Khim Metall Institu PROCESS FOR ENRICHING DIFFICULTLY ENRICHABLE OXIDE COPPER ORE
WO1988002408A1 (en) * 1986-07-29 1988-04-07 Khimiko-Metallurgichesky Institut Tsentralno-Kazak Method of concentration of difficult-to-concentrate oxidized copper ore
GB2204507B (en) * 1986-07-29 1990-06-27 Kazakhsh Khim Metall I Concentration of oxidized copper ores
DE3690783C2 (en) * 1986-10-04 1990-11-22 Shimiko Metallurg I Central Nc Oxide copper mineral concn. by flotation - involving pre-sulphidation with molten sulphur
US4735783A (en) * 1987-04-22 1988-04-05 Falconbridge Limited Process for increasing the selectivity of mineral flotation
US5443622A (en) * 1994-02-28 1995-08-22 Kennecott Corporation Hydrometallurgical processing of impurity streams generated during the pyrometallurgy of copper
US5616168A (en) * 1994-02-28 1997-04-01 Kennecott Utah Copper Corporation Hydrometallurgical processing of impurity streams generated during the pyrometallurgy of copper
WO1996025361A1 (en) * 1995-02-17 1996-08-22 Baker Hughes Incorporated Copper precipitation process
US20090241734A1 (en) * 2008-03-25 2009-10-01 Imagawa Harue Method of leaching copper sulfide ores containing chalcopyrite
AU2008249240B2 (en) * 2008-03-25 2011-02-10 Jx Nippon Mining & Metals Corporation Method of leaching copper sulfide ores containing chalcopyrite
WO2015081385A1 (en) * 2013-12-03 2015-06-11 The University Of Queensland Copper processing method
US20170191144A1 (en) * 2014-05-28 2017-07-06 Matthew Leslie Sutcliffe Method for ammoniacal leaching of copper from oxidised copper ores
US10590512B2 (en) * 2014-05-28 2020-03-17 Metaleach Limited Method for ammoniacal leaching of copper from oxidised copper ores

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JPS5117154B1 (en) 1976-05-31
AU3679271A (en) 1973-06-14
CA960466A (en) 1975-01-07
PH9284A (en) 1975-08-13
AU470987B2 (en) 1976-04-08
ZM18471A1 (en) 1973-10-22
ZA718277B (en) 1972-08-30

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