WO1996025361A1 - Copper precipitation process - Google Patents

Abstract

A method for separating copper and other metals in solution comprising the steps of precipitating the copper in a reactor at a free acid range of about 0.05 to 180 grams per liter, at a temperature from about 25 °C to about 90 °C in an aqueous solution with elemental sulfur, or chalcopyrite, and material selected from the group consisting of soluble sulfites and soluble bisulfites, and separating the precipitated copper, in the form of copper sulfides, by thickening the solution, recycling part to the precipitation step, and filtering copper sulfides from the other part.

Description

COPPER PRECIPITATION PROCESS

BACKGROUND OF THE INVENTION Field of the Invention: The present invention relates to the separation of copper from one or more metals. More particularly, the present invention relates to the precipitation of copper as a sulfide from aqueous acidic solutions while retaining other metals in the liquid phase for later precipitation as desired.

Description of Related Art: The need and desirability of precipitating a variety of metals from solution is well documented. The metals may be in solution in waste water, for instance, which must be discharged and dealt with in an environmentally responsible manner. Typical prior art methods for the removal of such metals are described in U.K. Patent Application GB 2216114A and in U.S. Patents 3,575,853; 3,617,559; 3,738,932; 4, 186,088; 4,465,597; 4,543, 189; 4,566,975; 4,606,829; 4,728,438; 4,572,882 and 5,039,428. It is desirable in certain circumstances to remove and separate copper from other metals which are in solution. It is further desirable to carry out such processes by means of readily available chemical compounds, at low process temperatures and pressures and at high reaction rates thereby minimizing the cost, process times and equipment size. Examples of such copper removal processes are disclosed in U.S. Patents 3,218, 161 ; 3,728,430 and 5,032, 175, Canadian Patent 1 ,020,363 and Society of Mining Engineers of AIME, Paper No. 80-73, entitled "The Electroslurry Process for Copper Recovery from Smelter and Refinery Wastes. "

Patent 3,218, 161 (Kunda et al.) gives a method for precipitating values of metals which form insoluble sulfides more readily than nickel in acid and neutral solutions having a pH from 1 to 7. The method is aimed at the selective precipitation of metals from nickel and cobalt at temperatures greater than 125 degrees Fahrenheit or 52 degrees Celsius (125°F or 52°C) by the use of sulfur dioxide and finely divided sulfur, the latter in excess of stoichiometric requirements. The major metal to be precipitated is copper and the pH is adjusted to be within pH 1 to 7 at a preferred temperature of 180°F (82°C). The method requires an inert, or substantially inert, atmosphere.

U.S. Patent 3,218, 161 claims a working range of pH 1 to pH 7 which is equivalent to a theoretical acidity of approximately 5 grams per liter sulfuric acid to 0.05 milligrams per liter sulfuric acid. The patent embraces solutions with acid strengths greater than the equivalent of pH 1 , e.g., sulfuric acid strengths greater than about 5 grams per liter, by claiming prior adjustment of the pH. In other words, if faced with an industrial solution with a pH of less than 1, the solution would need to be neutralized by an alkali to bring it within the stated range. Whereas waste waters are generally of low acidity, industrial process streams such as refinery electrolyte and acid plant blowdown are high in acidity, being typically 150 to 200 grams per liter sulfuric acid. However, in these cases, it may not be economical or desirable to partially neutralize the acidity of the solution to be purified so that it comes within a range of relatively low acidity. For example, if a bleed stream of refinery electrolyte could be purified without decreasing the acid content, it would be possible to return the purified electrolyte to the refinery. By this means, the need to add fresh acid and dispose of large quantities of neutralization product, whether as solids or liquids, would be avoided.

U.S. Patent 4,404,071 (Abe et al.) achieves precisely this outcome by the purification of copper refinery bleed using hydrogen sulfide as the precipitating reagent, in which case copper is precipitated but in combination with arsenic, antimony and bismuth. This precipitat is commonly returned to the copper smelting stage where the impurity elements, especially bismuth, are re-distributed into the products. If any of these impurities has reached a maximum permissible level in the product copper, the new intake of these impurities in concentrates and other new feeds will need to be controlled. It would be advantageous in the art, therefore, to provide a method wherein the copper in refinery bleed can be separated from other metals in the bleed.

Patent 3,728,430 (Clitheroe et al.) gives a method for the extraction of copper from ores by leaching with sulfites or bisulfites and the precipitation of the extracted copper from the leach solution with bisulfite, or sulfur dioxide, and elemental sulfur. A pH between 1 and 6 is preferred at a temperature between 80° and 212°F (27° to 100°C). The two operations can occur in the same reactor. The method provides a way of beneficiating a copper ore and the copper sulfide product is preferably floated away from the residual matter.

The hydrometallurgical extraction of copper from ore-bodies and concentrates, for example, is well known. In particular, a solution of sulfuric acid is arranged to percolate through metals-containing material to extract copper as copper sulfate from the copper minerals present. Copper minerals may be present as oxides, sulfides, oxysalts and the like, and may require the use of air, oxygen or bacteria to solubilize the copper. Typical equations describing these extractions are: • Cu2θ + 2 H₂SO₄ + 0.5 O2 → 2 CuSO₄ + 2 H₂O

• Cu₂S + 2.5 O + H2S . → 2 CuSO₄ + HzO

• CuCO3.Cu(OH)2 + 2.H2SO.1 → 2.CuSO₄ + CO2 + 3.H2O

The copper sulfate solution produced from the reaction may be processed to produce a valuable copper product in a number of ways. It can be chemically precipitated in various forms or, more generally, it may be electrowon to produce metallic copper in the form of a cathode. The electrowinning is usually performed after solvent extraction of the copper from the leach solution followed by transfer into a strong acid electrolyte. The solvent extraction step has the advantage of rejecting impurities in the copper sulfate solution and providing an optimum electrolyte for copper electrolysis.

Patent 3,728,430 elects to leach the oxidic, or mixed oxidic-sulfidic ore with sulfites rather than sulfuric acid, then precipitate the copper as a sulfide. A feature of the patent is the production of sulfuric acid by the copper precipitation process and the use of this excess acid to meet, to a greater or lesser extent, the natural acid demand of basic components in the oxide ore. However, the value of the copper sulfide precipitate is lower than cathode copper obtained by electrowinning the copper sulfate and many copper oxide ore-bodies are exploited by solvent extraction and electrowinning (SX-EW) to maximize the added value. It would be advantageous in the art, therefore, to provide a method whereby high added value copper could be produced and at the same time provide acid for the natural acid demand of the ore. An impediment to the universal application of SX-EW for ore-body extraction is the poor leach behavior of the common copper mineral, chalcopyrite, whether by sulfite or acid leach. It would be advantageous in the art, therefore, to provide a copper extraction process that facilitates the leaching of chalcopyrite and other copper minerals to produce an added value copper material. Canada Patent 1 ,020,363 (Hall et al.) provides a method for the separation of copper from arsenic, antimony or bismuth in aqueous solutions and slurries. The method uses sulfur and sulfur dioxide to precipitate copper sulfide from solutions that are adjusted to an acidity of less than pH 4 at a preferred temperature between 60° to 90°C. The method requires air to be essentially excluded. Canada Patent 1 ,020,363 claims a pH less than 4 to effect the separation of copper from other more acid soluble metal sulfides. The description does not give a range for acidity and, in effect, the method relates to slurries and solutions having sulfuric acid concentrations of 15 grams per liter (patent examples 1 and 2), equivalent to a theoretical pH of 0.5.

In investigating the use of sulfur and sulfur dioxide as precipitation agents for copper in sulfuric acid solutions, it is our experience that whereas low acidity (high pH) solutions ma have an acceptable precipitation rate, higher acid strength (low pH) solutions have a slow reaction rate. It is also our experience that the inception of the copper precipitation reaction i any acid strength solution is slow without a minimum level of solids in suspension. These undesirable features are not indicated in the prior patent art described above.

In the operation of metallurgical processes it is known for copper-bearing solid and liquid by-product streams, or in-process streams, to be produced. These streams often contai valuable levels of copper and other metals which make beneficial extraction of the metals worthwhile. It may also be necessary to selectively remove metals to benefit the overall metallurgical flowsheet and improve the economics of operation.

In one such application, it is known for copper smelters and associated refining operations to generate flue dusts, acid plant liquid bleeds and refinery liquid bleeds. The dust can be processed by hydrometallurgical means to produce a liquid that contains copper and other metals. In one instance, the liquid streams can be processed to extract copper and other metals by precipitation as sulfides using hydrogen sulfide gas or liquids such as sodium hydrosulfide. The cost of reagents based on hydrogen sulfide is high, however, and the quality of the resulting precipitates can be poor.

It would be advantageous in the art, therefore, to provide improved sulfide precipitation methods for copper over a wide range of acidity, which reduce the cost of reagents while providing a readily filtered precipitate at an acceptable precipitation rate. The ability to selectively extract copper would also be an advantage over the less selective nature of extraction using hydrogen sulfide-based reagents.

SUMMARY OF THE INVENTION In accordance with the present invention, copper is precipitated as a sulfide from aqueous acidic solutions, such as those produced from the processing of ore concentrates or non-ferrous smelter by-products, while retaining other metals in the liquid phase for optional downstream processing. The copper is precipitated over a wide range of acidity with relatively inexpensive reagents, at a desirable precipitation rate, and the copper precipitate is readily filterable. This invention also provides more economical means of precipitating copper by using less expensive and more readily available reagents.

As a broad statement of the present invention, copper sulfide is precipitated from an aqueous acidic solution containing soluble copper values by contacting the aqueous acidic solution with (i) sulfur, (ii) sulfur dioxide, and (iii) an added copper sulfide. The sulfur is usually supplied in the form of elemental sulfur or chalcopyrite, the sulfur dioxide is usually supplied as a soluble sulfite or bisulfite or as smelter gas, and the added copper sulfide as a thickened recycled stream of copper sulfide (which was originally precipitated as a product of the process). Sufficient copper sulfide is added to the process to ensure that the solids content of the precipitation zone (the zone in which the soluble copper is reacted with the precipitating sulfur reagents) exceeds the maximum solids content that would result in the zone if the precipitation reaction proceeded to completion (batch mode) or reached steady state (continuous mode) under the particular conditions of the precipitation of interest.

In one embodiment of this invention, copper is precipitated as a sulfide from an aqueous solution of copper and other metals, the solution having a free acid range of about 0.05 to about 200 grams per liter (gpl). The solution is contacted with elemental sulfur and a material selected from the group consisting of soluble sulfites or bisulfites, the contacting conducted at a temperature from about 40° to about 90°C. The precipitated copper sulfide is removed from the precipitation zone, thickened by any conventional technology, optionally filtered, and then a portion recycled to the precipitation zone.

A variable part of the thickened suspension is recycled to the precipitation stage to promote the reaction, increase the operable range of acidity, increase the rate of reaction, reduce the reaction time and enhance the degree of completion, all relative to a non-recycled operation. The precipitation may occur in a multicompartmental reactor, such as four compartments in series. Also, the method may use chalcopyrite as a source of sulfur or as a material to be processed in its own right.

BRIEF DESCRIPTION OF THE DRAWINGS FIG. 1 is a flow diagram of a first embodiment of the invention in which an aqueous solution of impure copper sulfate is processed using elemental sulfur and sulfur dioxide;

FIG. 2 is a flow diagram of a second embodiment of the invention in which a weak, impure solution of copper sulfate is processed into a strong, pure solution; FIG. 3 is a flow diagram of a third embodiment of the invention in which previous embodiments of FIGS. 1 and 2 are used to extract copper from oxidic copper to produce cathode copper;

FIG. 4 is a flow diagram of a fourth embodiment of the invention, similar to the previous embodiments of FIGS. 1 and 2, wherein elemental sulfur is substituted with sulfur derived from chalcopyrite-bearing ore;

FIG. 5 is a flow diagram of a fifth embodiment of the invention in which the embodiment illustrated if FIG. 4 is modified by routing a majority of the thickened solids and liquids to an oxidation stage; FIG. 6 is a flow diagram of a sixth embodiment of the invention in which the embodiment of FIG. 5 is modified to include a leach of oxidic copper ore;

FIG. 7 is a schematic diagram showing a seventh embodiment of the invention illustrating operational parameters and conditions;

FIG. 8 is a schematic diagram of a vessel used to simulate a reactor of the present invention;

FIG. 9 is a drawing of a laboratory simulation of the present invention shown in FIG. 7;

FIG. 10 is a correlation between rate and solids concentration in the reactors in Test 14 through 17 of the present invention; FIG. 11 is an illustration of separation of metals in each stage versus EMF for Test 14 of the present invention;

FIG. 12 is an illustration of separation of metals in each stage versus EMF for Test 16 of the present invention;

FIG. 13 is an illustration of copper concentration as a function of precipitation time of the present invention for Test 18 through 20;

FIGS. 14(a) and (b) are summaries of the sedimentation test on the precipitate produced in the process as set forth in Table 5 herein;

FIG. 15 is a summary of cake weight versus thickness from the filtration tests on the precipitate as set forth in Table 5 herein; FIG. 16 is a summary of cake weight versus formation time of the precipitate as set forth in Table 5 herein;

FIG. 17 is a summary of cake moisture versus dry time of the precipitate as set forth in Table 5 herein; FIG. 18 is a summary of wash time versus cake weight versus wash volume of the precipitate as set forth in Table 5 herein;

FIG. 19 is an illustration of residual copper concentration in solution as a function of precipitation time of the present invention using chalcopyrite with various EMF potentials;

FIGS. 20(a) and (b) illustrate settling data derived from thickening tests performed on precipitation slurry using chalcopyrite;

FIG. 21 is a graph illustrating test results of the inventive process conducted at high acidity levels; and

FIG. 22 is a schematic diagram of a generalized description of the invention.

DESCRIPTION OF THE INVENTION

FIG. 22 describes the process of this invention in a generic manner. Soluble copper (e.g. , copper sulfate, CuSOi) is contacted with elemental sulfur (S°) and sulfur dioxide (SO2) in a reaction vessel to precipitate the soluble copper as copper sulfide. The contacting is typically conducted with agitation (the stirrer) and at a temperature above ambient (e.g., in excess of 25°C). The soluble copper is usually associated with other soluble metals in an aqueous acidic solution with a free acid content in excess of 0.05 gpl.

Copper sulfide will continue to precipitate until the soluble copper or the precipitating reagents are depleted if the process is conducted in a batch mode, or until steady state is achieved if the process is conducted in a continuous mode. In either event, one hallmark of this invention is to increase the solids content, i.e., the copper sulfide, of the precipitation zone (here the reaction vessel) by adding copper sulfide. This addition has the desirable effect of increasing the precipitation rate.

The copper sulfide that is added to the reaction vessel is usually and preferably a thickened recycled stream. As the copper sulfide is formed in the reaction vessel, it is transferred to a thickening tank by any conventional means, e.g. , gravity overflow, pumping, etc. Here the copper sulfide content is thickened (in FIG. 22, from 30 gpl to 500 gpl). The thickened copper sulfide is then divided into two streams, a product stream for further processing (e.g., smelting or converting) and a recycle stream for return to the reaction vessel. Clear overflow is removed from the thickener as necessary to maintain a working balance of liquids and solids in the thickener. While FIG. 22 depicts the thickener as a separate vessel, the action of thickening the CuS can occur in the same vessel in which the CuS is precipitated. Vessels such as these will have a precipitation zone and a thickening zone and will allow for the transfer of thickened CuS from the latter to the former zone. The use of sulfur and SO2 as the precipitating reagents produces a precipitate that has desirable settling and filtering properties, and this is another hallmark of this invention. Moreover, this combination of precipitating reagents affords a more selective precipitation of copper than generally possible with the use of other precipitating a reagents, e.g., hydrogen sulfide (H2S). The conditions at which the process of this invention is operated will vary with the reagents, equipment, starting materials and the like. As such, operating conditions such as temperature, pressure, agitation, contact or residence time, molar ratios, solids content, acid strength, etc. , are selected to optimize copper precipitation in terms of rate, selectivity and completion. Our attempts to duplicate the teachings of the above patents concerning separation of a copper product, particularly at pH levels below 2 and at temperatures of 60°C have not been able to reach complete copper precipitation in a reasonable length of time (less than 2 hours) without employing the technique of increased solids content in the reactor. When we utilize the practical measure of recycling of thickened suspension to build solids content in a continuous operation, the reaction goes essentially to completion much more rapidly when compared at similar conditions to those cited in the above patents.

Furthermore, the practicality of the process is enhanced in the present invention by the application of smelter gas (SO2) containing appreciable quantities of free oxygen, whereas the prior art claims the substantial absence of oxygen. Whereas the prior art uses the chemistry of the method to separate copper from other metals by controlling the acidity of the system, the present invention provides for the sequential separation of metals by the control of the reduction potential, especially in those situations whereby copper, bismuth and antimony need to be sequentially separated.

The chemistry of the preferred embodiment of the present invention can be generally described and schematically represented as follows:

CuSO₄ + S + SO2 -I- 2 H2O → CuS + 2 H2SO.1 (1)

CuFeS₂ + 3 CuSO₄ + 2 SO2 + 4 H2O → 2 Cu₂S + 4 H2SO4 + FeSO₄ (2) The first reaction (1) above represents one embodiment of the invention which utilizes elemental sulfiir (S) and SO2 as reactants from suitable sources. The second reaction (2) above represents an alternative source of sulfur-bearing reagent in the preferred embodiment. In the description of the this invention, the term "copper sulfide" is used interchangeably for both the cupric and cuprous forms of this sulfide. Notably, the cuprous form of this sulfide may be formed as a secondary product during the precipitation reaction as described by equation (5). The precipitation reaction or step in all of the embodiments of this invention can be practiced using one or more, and preferably more, reaction vessels.

This first embodiment of the invention, shown schematically in FIG. 1 , is practiced by employing the impure aqueous sulfate solution from the natural (e.g. , mine run-off), or engineered (e.g., heap leaching), leaching process of a copper-containing material. Alternatively, a copper-containing aqueous sulfate stream from a production process (e.g. , refinery bleed) may be employed. The copper-containing aqueous sulfate streams are subjected to a precipitation process in a precipitation system. The solids suspension from the precipitation process is thickened and a portion of the thickened suspension is recycled to the precipitation stage as a slurry to initiate the reaction, increase the operable range of acidity, increase the rate of reaction, reduce the reaction time, and enhance the degree of completion, all relative to a non-recycled operation.

The product sulfide precipitate from the preferred embodiment may be separated from the thickened suspension by filtration and processed to final copper-containing product by other means, such as pyrometallurgical smelting methods. The residual filtrate solution (containing sulfuric acid, among others) from the reaction in this first case is either neutralized and discarded, or returned to the process for beneficial use (e.g., as described in FIG. 3). Alternatively, the residual filtrate solution which may contain additional valuable metals, e.g., molybdenum, cadmium, zinc, antimony, bismuth, arsenic, etc., may be processed to recover one or more of these metals.

In a second embodiment of this invention, shown schematically in FIG. 2, the product sulfide precipitate may be separated from the solution by filtration and re-oxidized to copper sulfate in an additional reactor system:

CΛ12S + 2.5 O2 -I- H₂SO₄ → 2 CuSO₄ + H2O (4) By this means, the copper can be selectively removed from impure solutions and concentrated in solution so that the resulting copper sulfate can be electrowon to copper cathode, or processed to other products such as copper sulfate crystals.

In a third embodiment of this invention, shown schematically in FIG. 3, the sulfuric acid solution remaining after removal of the copper sulfide precipitate may be utilized to provide acid for reaction with the basic minerals of an oxidic copper ore-body (e.g., an ore- body containing copper in a non-sulfidic form).

Oxidic copper ore-bodies may contain relatively high proportions of basic minerals that consume acid from circulating leach solutions used to convert the oxidized form of copper to soluble copper sulfate. Ore-bodies may consume inefficiently 10 to 20 kilograms (kg) of 100% sulfuric acid per ton of rock, and in extreme cases may consume as much as 100 kg of 100% sulfuric acid per ton of rock. This inefficiency has a large impact on the economics of the extraction process because processing sites are often in remote areas thereby necessitating the transport of acid to, or the generation of acid at, the site. In this third embodiment, sulfur may be brought to site to produce acid according to the preferred embodiment of the first reaction:

CuSO₄ + S + SO2 + 2 H2O → CuS + 2 H₂SO₄ (1)

It will be seen that for every weight unit of copper that reacts, 3.08 weight units of sulfuric acid are produced. On return to the copper extraction process, 1.54 of those weight units of sulfuric acid are consumed to produce a soluble weight unit of copper. Thus, 1.54 additional weight units of sulfuric acid are available from each copper extraction cycle to meet the consumption of acid by the basic components of the ore or concentrate.

At the limit, all of the copper will need to be precipitated as sulfide to supply the maximum quantity of acid to meet the inefficient consumption of ore or concentrate. Hence, for a common acid demand of 10 kg per ton of ore or concentrate, the extractable copper content will need to be 6.5 kg per ton, or 0.65 % copper in the material. This concentration of copper is common in ore-bodies. Hence it is seen that there is a relationship between the extractable copper present and the maximum acid demand of the rock. If the acid demand exceeds the limit, then additional acid will need to be brought in or generated on site, if it is economical to do so.

Under certain circumstances of this third embodiment, only a proportion of the extractable copper will need to be precipitated from solution to meet the acid demand of the rock. The remaining solution can be solvent extracted and electrolyzed in the conventional manner. Thus, the transport of acid from industrial centers to isolated areas and the high capital costs of on-site acid plants are both avoided. If the preferred product is cupric sulfide (CuS) the oxidation stage and the SX/EW stage are unnecessary. If the product sulfuric acid from the process exceeds the demand made by the basic ore, then the excess acid can be diverted to a waste water treatment operation.

If a second reactor system is provided in this third embodiment to the preferred reaction, the amount of sulfur brought to site can be reduced by using the cupric sulfide to react with additional copper sulfate and sulfur dioxide to produce cuprous sulfide:

CuSO4 + CuS + SO2 + 2 H2O → CU2S + 2 H:SO₄ (5) The same quantity of acid is still produced per weight unit of copper, but the sulfur quantity is reduced by 25% when taking into account the sulfur in the sulfur dioxide.

In circumstances whereby it is not favorable to make a copper sulfide product, the copper sulfides produced in the third embodiment may be oxidized according to the method expounded in the second embodiment, and the copper sulfate so produced routed to the SX/EW stage for production of cathode copper. As the SX/EW stage liberates an equivalent quantity of sulfuric acid from the input copper sulfate for return to the ore leaching step, an additional equivalent quantity of acid is available from the precipitation stage to meet the demands of the basic components of the ore-body.

In a fourth embodiment of the invention chalcopyritic-bearing ores or concentrates are processed:

CuFeS₂ + 3 CuSO₄ + 2 SO2 + 4 H2O → 2 Cu₂S + 4 H2SO4 -I- FeSO₄ (2)

In this fourth embodiment, a cuprous sulfide precipitate is produced that can be readily oxidized to copper sulfate according to the reaction:

2 CU2S + 5 O2 + 2 H2SO4 → 4 CuSO4 + 2 H2O (6) The cuprous sulfide precipitate may be removed from the solids suspension produced in reaction (2) by filtration or flotation.

The sulfuric acid used in the oxidation reaction may be at least partially derived from the precipitation reaction and as such will contain ferrous sulfate (FeSO ). Depending on the conditions of the oxidation reaction, some of the ferrous sulfate may be oxidized to ferric sulfate (Fe2(SO₄)₃). Since ferric ions can contaminate the SX/EW, the oxidation reaction is preferably conducted to disfavor the production of ferric sulfate. The sulfuric acid required in the oxidation of the cuprous sulfide by reaction (6) may be provided from the acid produced in reaction (2). According to the stoichiometry of the reactions, half of the solution produced in reaction (2) is required to meet the acid demand of reaction (6). The remaining half of the solution from reaction (2) is withdrawn as a bleed stream for iron and any other solubilized species. As most copper ores and concentrates contain basic minerals, such as limestone, dolomite and magnesite, the aqueous solution produced in reaction (2) may have lost some of the indicated sulfuric acid through neutralization of these minerals. The remainder may be treated with alkali to neutralize the excess acid and precipitate iron.

Three fourths of the copper sulfate solution produced by reaction (6) is returned to continue reaction (2) with new chalcopyrite. The other one fourth is removed and purified from soluble iron. The resulting copper sulfate may be electrowon to copper cathode, or processed to other products such as copper sulfate crystals. Sufficient sulfuric acid is produced in the SX stage to meet the acid demands of the copper oxidation reaction.

In a fifth embodiment of the invention, shown schematically on FIG. 5, chalcopyritic- bearing ores or concentrates are processed as in the fourth embodiment given above. However, in this fifth embodiment, the whole solution and solids suspension produced by reaction (2) is oxidized as follows: 2 C S + 4 H2SO4 + FeSO< + 5 V* O: → 4 CuSO- + V4 Fe:(Sθ4)3 + 2V4 H2O + VΛ H2SO4 (7) As previously noted, the quantity of acid available for this reaction will be affected by the basic materials present in the chalcopyritic-bearing ores or concentrates. The ferric sulfate that may be produced by this reaction is a strong oxidant and will assist in the oxidation of copper species in the concentrate or ore-body when three fourths of the copper sulfate solution is returned to continue reaction (2) with new chalcopyrite, as in the fourth embodiment above. The oxidation reaction product is separated into solid and liquid fractions using conventional technology. The solid fraction may be subjected to flotation to separate gangue material from a copper concentrate. This flotation operation is facilitated by the preceding precipitation and oxidation reactions. As in the fourth embodiment, one fourth of the above copper sulfate solution (i.e. the liquid fraction from the thickening operation) is removed and purified from soluble iron by solvent extraction. The resulting copper sulfate may be electrowon to copper cathode, or processed to other products such as copper sulfate crystals. In a sixth embodiment of the invention, shown schematically in FIG. 6, chalcopyritic- bearing ores or concentrates are processed as in the fourth and fifth embodiments given above. In this sixth embodiment, the free acid produced by reaction (2) in the fourth case and reaction (7) in the sixth case, is utilized for the purpose of reacting with oxidic copper material and acid-consuming basic components. These acid-consuming materials and components may be part of the chalcopyritic-bearing material, or separate oxidic and basic material. If the former, the material will be processed in the vessel. In the latter, the material will be processed on a heap.

On many occasions, oxide ore-bodies are found associated with sulfides that contain chalcopyrite. Hence, the economics of copper extraction may be improved by the above- described leaching of chalcopyrite-bearing ore, or concentrate, to provide acid for the basic components of the ore or concentrate. It may be only necessary to leach that quantity of chalcopyrite-bearing material needed to provide the required sulfuric acid, as the remaining sulfide minerals containing the chalcopyrite can be concentrated and transported to pyrometallurgical processing sites. Alternatively, copper concentrate may be extracted after the chalcopyrite consuming stage of the sixth embodiment.

By the above invention, ores and concentrates that contain chalcopyrite may be processed to cathode copper, at the mine site or in other locations, by simple hydrometallurgical operations. The major reagent to be supplied is sulfur dioxide which can be produced by the roasting of iron pyrite or the combustion of elemental sulfur. In a similar manner, oxidic ores can be processed to produce saleable copper sulfide concentrates.

The process of this invention is preferably conducted such that a thermal balance is maintained between the various operational steps. Each of the operational steps 13 conducted at conditions and in equipment that maximize their individual performance. The maintenance of the thermal balance between the various operational steps is accomplished using known and conventional techniques, i.e., heat exchangers.

The invention is further described by reference to FIGS. 7-21. As shown schematically in FIG. 7, a precipitation vessel 20 containing four reaction compartments 22 in series may be used. Alternatively, separate reactors may be used. Each reaction compartment 22 includes a turbine 24 for mixing, baffles 26, reagent sparger 28, suitable temperature and sample ports and level control system means. The aqueous solution containing copper is introduced into the precipitation vessel 20 and is subjected to the reagents therein while being transferred from one compartment 22 to the next. The discharge from the precipitation vessel 20 has flocculent added thereto and flows to the continuous thickener 30 where part of the underflow 31 therefrom is recirculated to the precipitation process and part (equivalent to the new copper fed to the process when in material balance) is fed to a filter 32 to produce a valuable filter cake. The overflow 33 from the thickener 30 is recirculated, in whole or in part, to the copper leaching process, or is passed forward, in whole or in part, to downstream systems 34 for removal of other metals. The liquid from these downstream systems 34 may also be returned, in whole or in part, to the copper leaching process.

One aspect of the embodiment shown in FIG. 7 is the processing of secondary copper materials such as flue dust, bag house dust, etc. These materials are leached in an acid media, e.g., sulfuric acid, to produce an impure solution containing copper sulfate and soluble species, e.g., bismuth, arsenic, antimony, cadmium, zinc, molybdenum, and iron, and a solid residue depleted of these metals. The product from this leach operation is separated into solid and liquid fractions. The solid fraction is routed for further processing, e.g. , recovery of precious metals. The liquid fraction is routed to a precipitation operation as shown in FIG. 1. In a manner similar to that of the first embodiment, copper sulfide is selectively precipitated from the impurities in the aqueous copper sulfate solution, more specifically arsenic, antimony, molybdenum, bismuth, cadmium, zinc and iron. The solution resulting from the separation of copper sulfide from the impurities is optionally in part recycled to provide acid for the dust leach operation, or is bled to further processing (which may consist of further removal of impurity elements for resale (e.g. , As) or an environmentally acceptable disposal, e.g., Bi or Sb).

Another feature of this embodiment is the introduction of copper-containing bleeds (e.g., acid plant blowdown (APB), refinery electrolytes, etc.) into the precipitation operation or stage to selectively precipitate the copper values. These bleeds may and usually do introduce additional acidity into the process and this enhanced acidity promotes selective precipitation of the copper values. Indeed, one hallmark of this invention is the ability to precipitate copper at the high acidity while promoting the solubility of the impurities, e.g., Bi and Sb.

The supply of sulfur dioxide in this embodiment may be provided, for example, from an associated smelter off-gas stream and in part from various bleed streams, e.g., APB, that contains soluble sulfur dioxide. The sulfur for the copper precipitation reaction may be provided, like in the other embodiments of this invention, from elemental sulfiir or chalcopyrite-containing concentrates from an associated smelter. FIG. 23 shows an alternative embodiment of FIG. 7 whereby the leach and precipitation operations are combined in a single reaction step. In this embodiment, copper- containing solids, e.g., electrostatic precipitation dusts, are fed to a reaction vessel optionally in conjunction with copper-containing aqueous bleeds. In this reaction vessel, the copper- containing solids are leached by sulfuric acid introduced as a separate reagent, or optionally recycled in an aqueous stream. The leachate so formed contains soluble copper, Bi, As, Sb, Cd and Fe. The soluble copper so formed is reacted in situ with sulfur dioxide and a source of sulf ir, such as elemental sulfur or chalcopyrite to form a copper sulfide precipitate. The aqueous suspension of copper sulfide and leach residue is removed from the reaction vessel to a thickening stage. In this stage the copper sulfide and leach residue are settled to form a thickened suspension, part of which is recycled to the reaction vessel. A portion of this thickened suspension is removed, optionally filtered, and routed for further processing, e.g. , through a copper smelter. The clear solution produced from the thickening stage is optionally partially recycled to the reaction vessel to utilize sulfuric acid and a portion is removed for further processing, e.g., removal of Bi, Sb, As, Cd and Fe, or to waste water treatment. A series of tests substantially embodying these conditions were run to simulate a continuous plant arrangement as generally described in FIG. 7. As shown in FIG. 8, to simulate the precipitation vessel 20 containing four reaction compartments 22, simulated reactors were used comprising four 2-liter glass beakers 36 each having an effective operating volume of 1.6 liters. Each simulated reactor was equipped with a Rushton turbine 38, baffles 40, reagent sparger 42, a temperature port 44 having a thermometer 46, an inlet 48 for feed liquid, and an outlet 50. Four such glass beaker reactors 36 were used in series to simulate the precipitation vessel 20 of FIG. 7.

The laboratory system is shown in its assembled arrangement in FIG. 9. Pumps 52 were used to provide a continuous flow of reactants, copper sulfate feed liquor, sulfur slurry, and sulfite solutions. Interstage transfer pumps 54 were also used to transfer slurry from one glass beaker reactor 36 to the next, and other pumps (not shown) were used to add flocculent to the final discharge and to recirculate the underflow from the continuous thickener. When sulfur dioxide gas was used, rotameters (not shown) were provided to monitor the flow of gas. To provide control of the acidity conditions in each reactor, a peristaltic pump was used to add sodium hydroxide to each of the reactors. Not shown in FIG. 9 is the continuous thickener. In general, the present invention is directed to the following preferred conditions using a reaction temperature of about 60°C and a solution pH corresponding to a free acid range of about 0.05 to about 180 gpl (grams per liter) H2SO4, although 10-20 gpl may be preferred. Approximately 10% excess sulfur may be added, although less may be required. Test 4, as referenced in Table 1 , shows that an excess amount of SO2, when supplied by NaHSO₃, is not required for the recycle ratio and acidity conditions employed. However, increased acidity, or reduced solids density, or shallow reactor vessel configuration may result in the necessity of adding an excess of SO2, supplied in gaseous form. That is, increased acidity will tend to decrease the solubility of SO2, reduced solids density will tend to decrease the adsorption of SO2, and shallow vessels will tend to decrease SO: contact time and bubble dispersion.

Furthermore, increased acidity will tend to oppose the production of copper sulfide according to the chemistry of the invention thereby requiring greater concentrations of SO2 to compensate. To overcome this potential limitation, SO2 may be introduced into the reaction vessels in a countercurrent manner to that of the copper-bearing solution, with the gas which exits from the last stage being returned to the second to the last stage and so on sequentially to the first stage. By this means, the efficiency of SO2 use can be maximized.

Solids precipitation substantially begins in the measured pH range of 2.7 to 3.0 for a feed liquor containing approximately 20 gpl copper as sulfate, 1 gpl arsenic, 10 gpl iron, and minor quantities of bismuth, cadmium, antimony and zinc. A solution containing the constituents as described was fed into the glass beaker reactors 36 at a rate of about 40 milliliters per minute (40 ml/min.), equivalent to 48 grams per hour (48 gm/hr) of copper or, in terms of equivalent copper sulfide, 72 gm/hr. Depending upon the flow conditions, other reactants, including recycled underflow, could add 15 to 40 ml/min. reactive volume such that the nominal retention time in the system could vary from 80 minutes to 116 minutes.

Elemental sulfur was provided having a particle size in the 10-20 micron range. An elemental sulfiir of the described particle size, designated "Super Fine", may be obtained from a process licensed to Innochem Engineering Company of Canada, such sulfur being produced by a process of injecting molten sulfur through a nozzle into a water bath. Alternately, flowers of sulfiir, a sulfur product from Geneva Steel, Provo, Utah, or sulfiir made from natural gas may be used. All types of elemental sulfur may require crushing and grinding before use. Sulfur dioxide sources may include sulfites and bisulfites, sulfur dioxide in a pure state or a blend of sulfur dioxide, nitrogen and oxygen, such as from smelter gas. Additionally, the sulfites and bisulfites may be derived from the sodium or calcium salts of sulfites and bisulfites. A first series of tests, numbered 1 through 7, employed sodium bisulfite at variable precipitate solids recycle ratios. The results of the tests using sodium bisulfate are summarized in Table 1. Table 1 represents the various test conditions with sodium bisulfite as the SO2 donor which were employed to evaluate the process in a continuous operating mode. Different types of sulfur and varying dosages of the reactants were utilized while employing a moderately high recycle ratio of precipitated solids to the reaction zone. This approach had been found to be very critical in batch tests in making the reaction proceed at a practical rate and with a high degree of completion. The continuous system confirmed the observations made from the batch test results.

Notably, in Test 6 shown in Table 1 , copper precipitation appeared to be lower due to higher copper contents in the reactors resulting from an extended down-time of the test reactors. With SO2 not being added in the final reactor (see footnote 4 to Table 1), the system was slow to come to equilibrium and the precipitation appeared low. If the test had been continued at the same conditions beyond the test time of 2.35 hours, equilibrium would have been reached and the results would be similar to those obtained in Test 7.

TABLE 1

PRECIPITATION OF COPPER USING ELEMENTAL SULFUR AND SODIUM BISULFITE

TEST RESULTS

Test Copper Balances (1) Sulfur SOj Seed Solids (2) Reactor (3) Average

Test Free

No. Duration Retention Time Acid

Concentration hrs gpl

In Out Percent Excess Excess Average Concentration Recycle

Hours gm/hr gm hr Precipitated Source gm/hr % gm hr % gpl Ratio

1 7.00 8.1 10.80 78.0 Superfine 27.0 11 84 73 161 11.7 1.12 35

2 7.30 48.7 2.30 95.0 Superfine 24.5 0 152 210 230 15 1.25 35

3 2.00 49.1 5.90 88.0 Superfine 28.9 17 98 98 260 16.9 1.26 35

00 4 3.00 49.1 2.80 94.0 Flowers of Sulfur 29.8 21 48 0 265 17.2 1.26 35

5 3.00 49.1 7.40 85.0 Flowers of Sulfur 29.8 21 71 44 270 17.6 1.26 35

6 2.35 47.9 12.70 73.5 Flowers of Sulfur 28.3 17 73 (4) 51 265 17.5 1.27 35

7 2.65 47.9 1.70 96.5 Flowers of Sulfur 28.3 17 77 (5) 60 261 17.3 1.27 35

(1) In: Reactor feed liquor, soluble copper Out: Thickener feed liquor, soluble copper

(2) Solids concentration in reactors;

Recycle ratio based on amount of solids recycled to maintain this concentration prior to precipilation, based on total amount of CuS that could be precipitated, if 100% were achieved.

(3) Total reactor live volume, 6,400 ml, divided by thickener feed rate, ml/hr.

(4) NaHSOi added to first three reactors only.

(5) NallSCh added to first two reactors only. Temperature: 60-63°C

A second series of tests, numbered 8 through 13, employed pure sulfur dioxide gas under conditions similar to those of the first series of tests. These test results are summarized in Table 2. These tests demonstrate the effectiveness of gaseous SO2 in the process, even under conditions not conducive to efficient gas absoφtion, the high acidity of the solution and the shallow liquid depth in the reactors.

These tests were conducted to confirm the acceptability of gaseous SO2 as a reductant without attempts to optimize its utilization. Typically, 100% excess gas gave high precipitation of copper when sufficient sulfur was present. From Test 10 on, the acid produced during the reaction was neutralized to an extent by the addition of NaOH. Precipitation was generally good under these conditions.

It is important to note that the required quantity of sulfur can be provided in the incoming copper solution and as long as it is present in sufficient excess in the final reactor, the reaction will proceed to a high level of completion if SO2 is present. On the other hand, the addition of SO2 in gaseous form has been demonstrated to be necessary to each and every stage to reach maximum completion. That is, the retention of gaseous SO2 between stages is low, especially in higher acidity solutions, such that the reaction is suppressed.

TABLE 2 PRECIPITATION OF COPPER USING ELEMENTAL SULFUR AND 100% SOJ GAS

TEST RESULTS

Test Copper Balances (1) Sulfur SO- Gas Seed Solids (2) Reactor (3) Average

Free (5) Duration Retention Time Acid

Test Concentration hrs No. gpi

In Out Percent Excess Excess Average Concentration Recycle

Hours gm hr gm hr Precipitated Source gm/hr % gm hr % gpl Ratio

8 3.50 47.9 13.50 72.0 Flowers of Sulfur 28.3 17 80 66 270 15.8 1.43 35

9 2.60 45.0 9.80 78.0 Flowers of Sulfur 24.7 9 105 132 290 17.9 1 .45 35

10 6.25 51.3 3.00 94.2 Flowers of Sulfur 28.4 10 105 104 300 16.2 1.45 28

1 1 4.67 52.0 0.03 99.9 Flowers of Sulfur 30.0 15 105 100 310 17.2 1.42 20 ro o 12 8.00 51.6 5.80 89.0 Flowers of Sulfur 25.₄ 0 105 102 315 16.9 1.45 40 j

13 4.75 52.0 0.04 99.9 Geneva Sulfur 33.0 26 105 100 418 22.1 1.48 12 j

(1) In: Reactor feed liquor, soluble copper Out: Thickener feed liquor, soluble copper

(2) Solids concentration in reactors;

Recycle ratio based on amount of solids recycled to maintain this concentration prior to precipilation, based on total amount of CuS that could be precipitated, if 100% were achieved.

(3) Total reactor live volume, 6,400 ml, divided by thickener feed rate, ml/hr.

(4) Temperature: 60-63 C

(5) Add NaOH to reactors to reduce amount of acid produced (runs 10-17) Add 50% of SO2 to first reactor, balance to others equally (runs 9-13)

The third series of tests confirmed the applicability of a simulated feed to a smelter acid plant, such as may apply from a copper flash smelting process. It was shown that despite the presence of a partial pressure of 5% oxygen in the simulated gas, there was little or no interference with the reductive power of the SO2. In addition, the level of excess SO2 was decreased to as low as 54% , and high precipitation efficiencies were confirmed. As previously described, it is not considered that this represents a lower limit to the SO2 requirement since several factors are involved in its utilization, not least of which is the shallowness of the test reactors.

This third series of tests, numbered 14 through 17, utilized sulfur dioxide derived from a simulated flash smelter gas containing 35 % SO2, 60% N2 and 5 % 02. In such a flash smelting process, the smelting operation is continuous and the off-gas which passes to an acid plant contains a high level of sulfur dioxide with a relatively low oxygen content. The gas composition is also fairly constant due to the continuous nature of the process. Other continuous processes like the Noranda, Mitsubishi, and Isasmelt Processes produce high strength off-gases that may prove acceptable as a source of SO2 reductant. The recycle ratios for precipitated solids were varied from 0 to as high as 23, with the reaction times decreasing proportionately. These results are summarized in Table 3. The similarity in these results and those with pure sulfur dioxide (Table 2) indicates that this approach to copper precipitation would be very practical in that a process gas stream containing sulfur dioxide gas, could be employed just as effectively as a purified (and more costly) reactant, such as liquefied SO2 or sulfites/bisulfites.

TABLE 3 PRECIPITATION OF COPPER USING ELEMENTAL SULFUR AND SMFLTER GAS

TEST RESULTS

Test Copper Balances (1) Sulfur SOJ Seed Solids (2) Reactor (3) Average

Test Duration Retention Time Acidity

No hrs gpi

In Out Percent Excess Excess Average Concentration Recycle

Hours gm hr gm/hr Precipitated Source gm/hr % gm/hr % gpl Ratio

14 3 4 51 3 0 025 99 9 Geneva Sulfur 27 5 6 105 103 487 22 0 1 42 12

15 3 0 52 0 1 120 97 8 Superfine 31 5 20 84 60 165 10 6 1 30 6

16 3 0 54 0 1 690 96 9 Superfine 32 9 21 84 54 124 5 2 1 67 9

17 5 0 54 2 1 950 96 4 Geneva & Superfine 31 8 17 84 54 23 (4) 0 0 1 92 11 ro ro

(1) In Reactor feed liquor, soluble copper Oul Thickener feed liquor, soluble copper

(2) Solids concentration in reactors.

Recycle ratio based on amount of solids recycled lo maintain this concentration prior to precipitation based on lotal amount of CuS that could be precipitated, if 100% were achieved

(3) Total reactor live volume, 6,400 ml, divided by thickener feed rate, ml/hr

(4) Each reactor pre dosed with amount of precipitate corresponding to — 2596 precipitation occurring in each Temperature 60 63°C

The final test in this third series, Test 17, was designed to verify the results when using no recycled solids, which in previous studies had shown greatly reduced precipitation rates. It was assumed that there could be some precipitation of copper sulfide, and at best, this might be 25 % of the total quantity in each reactor, such that the last reactor would contain approximately 30 gpl sulfide solids if all the copper had precipitated. The reactors were prepared prior to the beginning of the tests so that these possible equilibrium conditions would exist initially by dosing with precipitated solids as if this degree of precipitation had occurred. Caustic soda was then added during the run at the dose required to maintain constant acidity assuming the precipitation would occur at a constant rate. The results were initially unexpected, when compared to the original batch test work, in that the final effluent from the system was almost as good as that with a higher seed precipitated solids concentration. The initial precipitation rate was very low, as expected, but in the second and third reactors the rate accelerated greatly, actually exceeding the rates observed in the other tests. Upon consideration, however, it appeared that the reaction rate was due to the reduced acidity since caustic soda had been added earlier before any acid would have been generated, thereby lowering the acid level to about 1 to 2 gpl by the time the precipitation rate began to accelerate. Based on the earlier batch test, it is reasonable to assume that, if seed solids had not been added nor had sodium hydroxide been added to neutralize an assumed generation of acid, there would have been very little precipitation occurring and the final copper concentration in the liquor proceeding to the thickener would have been close to that in the feed liquor.

A correlation between rate and solids concentration is best seen in FIG. 10 showing the copper concentrations in each of the reactors for the four runs, as well as the calculated instantaneous concentration as the feed liquor and recycled underflow entered the first reactor. The samples were taken from each reactor and, therefore would represent the concentration of copper leaving that reactor stage and entering the next reactor. These data are used to calculate the reaction rates shown in Table 4. TABLE 4

SUMMARY OF COPPER PRECIPITATION RATE DATA

TEST RESULTS

* Average values; acidity and reaction rates varied during tests

A good separation of copper from the other constituents of the solution is possible, provided the electromotive force (EMF) does not drop too low. This is best illustrated in FIGS. 5 and 6, representing the analytical results from Tests 14 and 16. In Test 14, where the copper sulfide concentration was 490 gpl, the reaction was complete after only 40 minutes of retention time, or as the slurry left the second stage reactor. It will be noted that bismuth began to precipitate in the second reactor and was completely precipitated in the third reactor. Antimony began to precipitate in the third reactor and it was almost completely removed from the solution leaving the fourth reactor. The EMF had dropped below 200 millivolts in the third stage and had dropped to 65 in the fourth. It had been indicated that the EMF remains around 250-300 millivolts during most of the precipitation reaction and then drops down to close to 200 millivolts as most of the copper is depleted. This provides a good control method for a continuous circuit, modulating the rate of liquor or sulfur dioxide addition in order to maintain the final EMF around 200 millivolts. It should be noted that these EMF measure¬ ments were taken on solutions that had been filtered and cooled, in order to obtain values that could be compared.

Test 16, which would represent a desirable operating level in terms of good separation of copper from the other elements, illustrates these preferred conditions. The EMF in this case was slightly lower than that noted in the previous test, although this may have been due to a difference in the sulfur dioxide concentration in the different samples. To determine if the system would be effective at higher acidities, Tests 18, 19 and 20, as shown in Table 4, were carried out on a batch basis, but with the simulated smelter gas as the source of sulfur dioxide. Since the process solution could be present at a higher temperature in a flue dust leach process, these tests were carried out at 80°C which would be anticipated to increase the reaction rate. The amount of precipitate solids present in the reactor corresponded to that which would exist with the recycle ratio of 23. In Tests 18 and 19, the acidity was allowed to increase due to the reaction, with the average acidity used in evaluating the tests. In Test 20, the acidity was controlled at a constant level averaging 37 gpl by adding sodium hydroxide during the course of the test. These results are represented in FIG. 13 which shows the copper concentration as a function of precipitation time. When the acidity is held constant, as in Test 20, the precipitation rate is virtually constant, and the reaction is complete in about 40 minutes. Precipitation rates calculated from these data are also shown in Table 4. It should be noted from Figure 5 that the major impurities, iron and arsenic, remained in solution while all of the copper was precipitated. Of the minor impurities, zinc and cadmium also remained in solution. Only bismuth and antimony appear to be all or partially precipitated, and this begins after approximately 80% of the copper has been precipitated.

The material produced in the inventive process was more easily thickened and filtered than that, for example, resulting from the use of hydrogen sulfide precipitant. A settling test and several filtration tests were carried out on the material produced during the continuous precipitation study, with the test conditions and results shown in Tables 5 and 6, and FIGS. 8 through 12.

TABLE 5

Flocculent: PERCOL 351 from Allied Colloids

Concentration (gpl): 1.00

Volume Added (ml): 2.00

Dosage (kilograms per metric ton): 0.0049

Vol (ml) Weight (g) Tare (g)

Undecanted Slurry: 2000.00 2996.00 489.00

Decanted Slurry: 280.00 1114.60

Dry Solids: 407.50 0.00

Settling Vessel Size (milliliters per centimeter): 47.05

Ultimate Interface Height ( illiliter per meter): 886.00

Sp. Grav. Supernatant: 1.0940

Sp. Grav. Solids: 5.0500

Depth Correction: AUTOMATED

Initial Final Ultimate

Minimum Desired Underflow Percent Solids: 40.00 16.25 64.57 66.30

UF Concentration (kilograms per liter): 0.05 0.02 0.12 0.13

Tm(min) Ht(ml) Tm(min) Ht(ml)

0.00 2000.0 9.00 310.0

0.50 1 105.0 10.00 302.0

1.00 630.0 12.00 295.0

2.00 470.0 15.00 290.0

3.00 430.0 20.00 285.0

4.00 395.0 0.00 0.0

5.00 370.0 0.00 0.0

6.00 350.0 0.00 0.0

7.00 335.0 0.00 0.0

8.00 322.0 0.00 0.0 Underflow Underflow Test Results Scale-Up

Weight Copper Unit Area, Unit Area,

Percent (Kilograms Square Meter Square Meter Per

Solids Per Liter) Per Metric Metric Ton Per Day

Ton Per Day (Depth-Corrected)

64.57 0.1194 0.1581 0.0395

64.00 0.1173 0.1248 0.0319

63.00 0.1136 0.0915 0.0235

62.00 0.1101 0.0716 0.0192

61.00 0.1067 0.0652 0.0170

60.00 0.1034 0.0551 0.0155

59.00 0.1002 0.0473 0.0146

58.00 0.0971 0.0453 0.0143

57.00 0.0941 0.0437 0.0142

56.00 0.0911 0.0394 0.0142

55.00 0.0883 0.0355 0.0142

54.00 0.0855 0.0320 0.0141

53.00 0.0828 0.0288 0.0138

52.00 0.0802 0.0259 0.0135

51.00 0.0776 0.0232 0.0131

50.00 0.0751 0.0207 0.0123

49.00 0.0727 0.0186 0.0117

TABLE 6

Material: COPPER SULFIDE PRECIPITATE Filter Cloth: POPR 929

Feed Tot Solids (weight percent): 66.000 Liquid Sp. Gr: 1.0900 Feed Diss Solids (weight percent): 0.000 Slurry Sp. Gr: 2.2600 Calcd Feed Susp SIds (weight percent): 66.000 Calcd Solid Sp. Gr.: 5.0555

Filter Area (square meter): 0.09 Barometric pressure (centimeters Hg): 65.354 Shiny pH: 2.0 Filtrate Susp Solids (milograms per liter): N/A

Slurry Temp (Decree C): 50.0 Air Leakage (square meter per minute): N/A Sluπy Feed Technique: TOP Air Flow Meter: GAS METER

Teat No. 1 2 3 4

Form Vacuum (centimeters Hg) 38.100 38.100 38.100 38.100

Wash Vacuum (centimeters Hg) 38.100 38.100 38.100 38.100

Dry Vacuum (centimeters Hg) 38.100 38.100 38.100 38.100

Form Tune (seconds) 2.000 3.000 4.000 2.000

Wash Time (seconds) 8.000 10.000 12.000 8.000

Dry Time (seconds) 60.000 45.000 30.000 45.000

Air Reading Bef (cubic meters) 0.000 0.000 0.000 0.000

Air Reading Aft (cubic meters) 0.000 0.000 0.000 0.000

Filter Vol (incl waβh)(milliliιer) 160.000 169.000 168.000 220.000

Wash Volume (milliliter) 100.000 100.000 100.000 100.000

Cake Thickness (millimeter) 13.000 15.000 17.000 22.000

Cake Tare Weight (gram) 7.170 7.100 7.220 7.180 Cake Total Wet Weight (gram) 226.480 268.200 312.690 432.160

Cake Partial Wet Weight (gram) 0.000 0.000 0.000 0.000

Cake Partial Dry Weight (gram) 191.810 226.580 260.780 382.260

Net Dry Cake Wt (gram) 184.640 219.480 253.560 375.080

Cake Moisture (weight percent) 15.809 15.940 16.993 11.742

Air Flow (cubic meters per minute 0.000 0.000 0.000 0.000 per square meter)

Bk Calc Feed S.S. (weight percent) 64.852 65.261 66.798 67.487

Cake Loading (Kilograms Per Meiers 0.199 0.236 0.273 0.403

Squared)

Moist Factor (minute-square meter 5.039 3.175 1.844 1.864 per kilogram)

Moist Factor 2 (graph 4) 0.000 0.000 0.000 0.000

Wash Displacements 2.884 2.403 1.926 2.004

WVw (kilograms per square meter) 23.83 28.34 3₂.75 48.44

Oilers per square meter)

Graph Symbol STAR STAR STAR PLUS

THICKENER CYCLONE UNDERFLOW UNDERFLOW

It should be noted that the quantity of recycled precipitate solids appears to be less important in maintaining a high reaction rate if the measured pH is maintained close to 2. However, high recycle ratios that may be used to maintain a productive reaction rate incidentally will also produce a substantially coarser particle that will yield lower cake moisture and give higher filtration rates on the product slurry. If the system is to be used following a pH reduction stage (e.g. , a bismuth precipitation stage), then it may be advantageous to add caustic soda or other suitable alkaline material to maintain the pH constant since this will maximize the precipitation rate. However, if this reaction were to be employed directly after a metals solubilization stage, it may be preferable to allow the acidity to increase during the precipitation reaction so that all, or a portion, of the resulting effluent could be recycled back to the leaching process as the lixiviant. This is especially true in the case of leaching ores and concentrates where the acid generated can be usefully employed in meeting the natural acid demand of the solids being leached. As an alternative to the use of elemental sulfur, chalcopyrite (CuFeS∑) may be used. A test was run under the conditions previously described using chalcopyrite as a substitute. The chalcopyrite used in the test was obtained from sources located at Bingham Canyon, Utah. Test 21, shown in FIG. 19, illustrates the test results. A sample of 250 grams of chalcopyrite was rod-milled for about 30 minutes to reduce the average particle size, and this was employed in the test. The reaction rate curve as shown in Test 21 was somewhat different from the results using sulfur, since there appeared to be a lag time before the reaction commenced, although subsequently the rate was relatively high. The test was also conducted at 80°C.

A thickening test was run on the slurry produced, the results being shown in FIGS. 14(a) and (b). Notably, the upper line in FIG. 20(b) represents the settling rate values calculated from the settling test data illustrated in FIG. 20(a), while the lower line represents a scale-up application to a full-sized thickener of the values calculated in the upper line. The underflow from this test was filtered at a rate/ of about 2.53 kilograms per hour per square meter (60 pounds per hour per square foot) yielding a filter cake having a solids concentration of 74.6% . An alternative approach to grinding chalcopyrite for use in the precipitation process described above is to process an amount of chalcopyrite through a cyclone of a diameter suitable to remove only the fine particles in the slurry — for example, the particles sized at equal to or less than 15 microns. This material may be thickened and added to the reactors. To summarize, the test work indicated that a retention time of about 80 minutes in a continuous circuit with a recycled solids concentration in the reactors of 300-400 gpl copper sulfide solids would yield almost 100% recovery of the copper and at the same time provide a good separation from the impurities, arsenic, iron, and other minor elements. The tests show that all the sulfur and sulfur dioxide sources tested are equally applicable. Sulfur dioxide at the concentration present in the smelter gas reacted as rapidly as did pure sulfur dioxide gas. An excess dose of 50-100% sulfur dioxide was required due in part to the shallowness of the test reactors and, in part, to varying acidity of the solutions. Approximately 10% excess elemental sulfur is sufficient to ensure reaction completion.

The solution acidity was found to be a major factor in controlling the rate of reaction. During precipitation, approximately 3 gpl H₂SO₄ was produced for every gram per liter of copper that was precipitated using sulfur dioxide. The reaction rate in a solution at 60°C and an initial pH of about 2 will decrease from approximately 10 kilograms copper per cubic meter per hour to about half that rate as the reaction progresses. To compensate for this, caustic soda or other suitable alkali can be added to maintain acidity at a constant level to yield a relatively constant precipitation rate. It should be noted that the use of NaHSO3 to provide the necessary SCh for the reaction, also results in a simultaneous neutralization of the acidic solution to help maintain acidity at a constant level. Sodium bisulfite can be obtained commercially or by reaction of SO2 with caustic soda.

A soluble neutralization product, such as is obtained by use of caustic soda or sodium bisulfite, is preferred to prevent the dilution of the copper sulfide by insoluble products of neutralization. Under these conditions, the effect of recycled solids concentration on the reaction rate is less pronounced, and the reaction will go to completion at concentrations as low as 20-40 gpl copper sulfide. A high concentration, in the range of 500 gpl copper sulfide, effectively doubles the reaction rate. Higher concentrations also produce a precipitate with a substantially coarser particle size. To simulate the use of the process on liquor at a much higher acidity, tests were conducted at 80°C and with a recycled solids concentration of 500 gpl. The average acidity varied from approximately 125 gpl to 40 gpl of HiSOt acid resulting in the reaction rate increasing from 5 to 30 kilograms copper per cubic meter per hour. This concentration of solids requires a recycle ratio of about 20 times the amount of solids being precipitated on each pass.

To determine the efficacy of the present invention at high acidity levels such as those likely to be encountered in electrolyte, for example in copper tankhouse circuits, Tests 22 and 23 were conducted. In these tests the acidity of a solution containing a final 20 grams per liter copper as copper sulfate was adjusted to 180 grams per liter sulfuric acid prior to reaction at 80°C.

In both tests, an excess of elemental sulfur was added just before commencement of the batch reaction. Sodium bisulfite was added as a 30% solution as a sulfur dioxide donor, with sulfuric acid being added to compensate for the dilution and neutralizing effects of this addition. By this means, the sulfuric acid level throughout the two tests was maintained around the 180 grams per liter level. Test 22 had no solids apart from the elemental sulfur added to the solution whereas Test 23 had 500 grams per liter of copper sulfide precipitate added to the solution.

Results from the two tests are shown on FIG. 21. In Test 22 with no added copper sulfide, there was only a small decrease in the copper concentration of the solution over a period of 4 hours. In Test 23 there was almost a complete precipitation of copper in 2 hours demonstrating the benefits of the present invention. The precipitate resulting from the process was easily thickened and filtered, with a thickener unit area in the range of 0.015 square meter per metric ton per day (0.15 square feet per ton per day) and filtration rates varying from 5.06 to 12.64 kilograms per hour per square meter (120 to 300 pounds per hour per square feet). Filter cake moistures ranged from 25% to as low as 12% when a simulated cyclone underflow was dewatered.

These high acidity tests show that not only do iron, cadmium, arsenic and zinc remained in solution while copper sulfide is precipitated, but that Bi and Sb actually increased in soluble concentration (as a result of being leached from the seed material used in these experiments). This illustrates that increasing the concentration of the acidity of the copper precipitating solution facilitates the selective precipitation of copper (which is particularly useful when operating with materials that contain many other metals, e.g. , APB).

In a method as described, the preferred temperature of the precipitation reaction is between 60° and 80°C whereas the temperature of the leach solutions may be in the range 5° to 80°C depending on source. The heat required for the precipitation reaction can be provided by recuperation of the heat content in the outgoing stream with the incoming stream, or the sensible heat of the SO₂-bearing gas stream or both.

Claims

CLAIMS What is claimed is:

1. A method for precipitating copper sulfide from an acidic aqueous solution containing soluble copper values, the method comprising contacting the solution with sulfur, sulfur dioxide and added solid copper sulfide.

2. The method of claim 1 in which the solution has a free acid content of at least about 0.05 gpl.

3. The method of claim 2 in which the free acid is sulfuric acid.

4. The method of claim 3 in which

A. The soluble copper values are precipitated as copper sulfide in a precipitation zone, B. The precipitated copper sulfide is recovered from the precipitation zone, and C. The recovered precipitated copper sulfide is thickened.

5. The method of claim 4 in which at least a portion of the thickened recovered precipitated copper sulfide is recycled as the added solid copper sulfide.

6. The method of claim 5 in which the solution is contacted with the sulfur, sulfur dioxide and added solid copper sulfide at a temperature of at least about 40°C.

7. The method of claim 1 in which the sulfur is elemental sulfur.

8. The method of claim 7 in which the elemental sulfur is added in a stoichiometric excess to that needed to precipitate the soluble copper values.

9. The method of claim 8 in which the elemental sulfur is present in a stoichiometric excess of at least about 10% .

10. The method of claim 5 in which the sulfur dioxide is formed in situ in the precipitation zone from a material selected from the group consisting of soluble sulfites and soluble bisulfites.

11. The method of claim 5 in which the sulfur dioxide is fed to the precipitation zone in the form of smelter gas.

12. The method of claim 5 in which the sulfur is provided in the form of chalcopyrite-bearing solids.

13. The method of claim 12 in which the chalcopyrite-bearing solids are added in a stoichiometric excess to that needed to precipitate the soluble copper values.

14. The method of claim 13 in which the chalcopyrite-bearing solids are present in a stoichiometric excess of at least about 10% .

15. The method of claim 5 in which the sulfur is provided as both elemental sulfur and in the form of chalcopyrite-bearing solids.

16. The method of claim 15 in which the elemental sulfur and the sulfur provided in the form of chalcopyrite-bearing solids are contacted with the aqueous solution at different points in the precipitation process.

17. The method of claim 12 in which the sulfur dioxide is formed in situ in the precipitation zone from a material selected from the group consisting of soluble sulfites and soluble bisulfites.

18. The method of claim 12 in which the sulfur dioxide is fed to the precipitation zone in the form of smelter gas.

19. The method of claim 5 comprising the further steps of: A. Oxidizing at least a portion of the thickened copper sulfide to produce a solution of copper sulfate; and

B. Electrowinning the copper sulfate to produce cathode copper.

20. The method of claim 19 comprising the further step of purifying the solution containing copper sulfate.

21. The method of claim 20 in which the solution containing copper sulfate is purified by solvent extraction.

22. The method of claim 19 comprising the further steps of:

A. Recovering sulfuric acid from at least one of the precipitation and electrowinning steps;

B. Introducing the sulfuric acid to a copper oxide ore body; C. Leaching the copper ore-body with the sulfuric acid to form copper sulfate in solution, the solution containing copper sulfate at a free acid range of at least about 0.05 gpl;

D. Precipitating copper from the solution as copper sulfide by the addition of sulfur and sulfur dioxide, and adding a thickened suspension to the solution to form a suspension of copper sulfide; and

E. Recycling sulfuric acid to the copper oxide ore-body.

23. The method of claim 22 comprising the further steps of:

A. Solvent extracting and electrowinning a portion of the solution of coppe sulfate to form copper cathode and a residual solution containing sulfuri acid;

B. Recycling the residual solution to the oxide ore-body.

24. The method of claim 23 in which smelter gas containing about thirty-five percent sulfiir dioxide and about five percent free oxygen is added to the aqueous solution during the precipitation to provide the sulfur dioxide.

25. The method of claim 24 in which excess smelter gas remaining from addition to the aqueous solution is recycled to a smelter acid plant gas stream.

26. The method of claim 5 in which the sulfur dioxide is contacted with the aqueous acid solution in a countercurrent manner.

27. The method of claim 5 comprising the further step of recovering soluble metals other than copper from the clear solution formed during the thickening step.

28. The method of claim 5 in which the precipitating of copper values from the aqueous solution is conducted at an electromotive force at or above 200 millivolts to selectively precipitate copper in preference to other metals in solution.

29. The method of claim 28 comprising the further step of continuing to process the aqueous solution after precipitate of the copper by the continuing addition of sulfur and sulfiir dioxide to precipitate other metals in solution.

30. The method of claim 29 in which the other metals precipitated from the aqueous solution include bismuth and antimony.

31. A method of separating copper and other metals from an aqueous solution, the method comprising the steps of:

A. Contacting an aqueous suspension containing copper- and other metal— bearing solids and sulfuric acid, the solution having a free acid range of at least about 0.05 gpl, with sulfur and sulfur dioxide for a time sufficient to precipitate copper as copper sulfide;

B. Recovering a suspension of the copper sulfide and adding to it a thickening agent to form a thickened suspension of copper sulfide; and C. Recycling at least a portion of the thickened suspension to the aqueous suspension to increase its solids content thereby increasing the rate of precipitation of the copper sulfide.

32. The method of claim 31 in which the solids content is increased to about 550 gpl to enable the use of a free acid level in the precipitation step of about 180 gpl free sulfuric acid.

33. The method of claim 32 in which the aqueous suspension is contacted with the sulfur and sulfur dioxide for a period of time of up to about two hours.

34. The method of claim 33 in which the copper is selectively precipitated from the aqueous suspension by maintaining the oxidation/reduction potential of the aqueous suspension at about 200 millivolts.

35. The method of claim 34 in which the sulfur is provided in the form of elemental sulfiir.

36. The method of claim 34 in which the sulfur is provided in the form of chalcopyrite-containing solids.

37. The method of claim 35 in which the sulfur dioxide is formed in situ from a material selected from the group consisting of soluble sulfite and soluble bisulfite.

38. The method of claim 35 in which the sulfur dioxide is formed in situ from a material selected from the group consisting of the sodium salts of soluble sulfites and soluble bisulfites, and calcium salts of soluble sulfites and soluble bisulfites.

39. The method of claim 38 in which step A is conducted in a sequential series of reactor vessels.

40. The method of claim 39 in which the sulfur dioxide is provided in the form of smelter gas.

41. The method of claim 40 comprising the further steps of:

C. Processing the aqueous suspension through the sequential series of reactor vessels in sequential stages of selective precipitation of copper until the copper is substantially precipitated from the aqueous suspension forming a spent processed aqueous suspension; and

D. Recycling a portion of the spent processed aqueous suspension to the first reactor vessel in series.

42. The method of claim 41 in which the aqueous suspension is derived at least in part from nonferrous smelter refinery bleed streams.

43. A method for precipitating cuprous sulfide from an aqueous solution containing soluble copper values and sulfuric acid, the method comprising the steps of:

A. Contacting the aqueous solution with sulfur dioxide, sulfur and a thickened suspension of copper sulfide, the solution having a free acid concentration of at least about 0.05 gpl, to precipitate cupric sulfide; B. Contacting the precipitated cupric sulfide with additional sulfur dioxide and thickened copper sulfide to precipitate a solids suspension of cuprous sulfide; C. Thickening the precipitated cuprous sulfide into a solids slurry and a clear overflow; and D. Recycling a portion of the solids slurry of step C for contacting with the solids suspension of step A and step B.

44. The method of claim 43 comprising the further step of thickening the cupric sulfide produced in step A and recycling a portion of thickened cupric sulfide to step A.

45. The method of claim 44 comprising the further steps of: A. Oxidizing the recovered cuprous sulfide to copper sulfate; and

B. Electrowinning said copper sulfate to produce cathode copper.

46. A method for precipitating copper sulfide from an acidic aqueous solution obtained from leaching secondary copper-bearing materials in which the copper is preferentially separated from other metals in the materials, the method comprising contacting the solution with sulfur, sulfur dioxide and added solid copper sulfide.

47. The method of claim 46 in which the solution has a free acid content of at least about 0.05 gpl.

48. The method of claim 47 in which the free acid is sulfuric acid.

49. The method of claim 48 in which A. The copper is precipitated as copper sulfide in a precipitation zone,

B. The precipitated copper sulfide is recovered from the precipitation zone, and

C. The recovered precipitated copper sulfide is thickened.

50. The method of claim 49 in which at least a portion of the thickened recovered precipitated copper sulfide is recycled as the added solid copper sulfide.

51. The method of claim 50 in which the secondary copper-bearing material is smelter dust.

52. The process of claim 51 in which the sulfur dioxide is fed into the precipitation zone in the form of smelter gas.

53. The process of claim 52 in which the sulfur is elemental sulfur.

54. The method of claim 53 in which the sulfur is provided in the precipitation zone from chalcopyrite-bearing solids.

55. Copper sulfide produced by the method of claim 1.

56. A method of producing copper sulfide from an aqueous solution containing chalcopyritic-bearing solids comprising the steps of:

A. Contacting an aqueous solution containing chalcopyrite with sulfiir dioxide and a solution of copper sulfate to produce a solid suspension of cuprous sulfide and leach residue;

B. Thickening the solid suspension to produce a thickened suspension of cuprous sulfide and leach residue and a clear solution containing sulfuric acid;

C. Recycling a portion of the thickened suspension of cuprous sulfide and leach solution to step A; and

D. Removing the remaining thickened suspension of cuprous sulfide and leach residue from step C for further processing.

57. The method of claim 56 wherein the remaining thickened suspension from step D is further processed comprising the steps of:

E. Contacting the thickened suspension of cuprous sulfide and leach residue with oxygen to produce copper sulfate in solution;

F. Removing residual solids from the copper sulfate solution from step E for further processing; G. Recycling a portion of the copper sulfate solution from step F to step A;

H. Removing the remaining portion of copper sulfate solution from step F for further processing by solvent extraction/electrowinning to produce cathode copper and a copper-containing raffinate; and I. Further processing the raffinate from step H.

58. The method of claim 57 comprising the further step of processing at least a portion of the raffinate from step H by returning said portion to step E.

59. The method of claim 58 comprising the further step of processing the remainin portion of raffinate from step H by routing to waste water treatment.

60. The method of claim 57 comprising the further step of processing at least a portion of the raffinate from step H by routing said portion to waste water treatment.

61. The method of claim 56 comprising the further steps of:

J. Processing the thickened suspension of cuprous sulfide and leach residu in a flotation step to produce a cuprous sulfide concentrate and a leach residue; K. Processing the cuprous sulfide concentrate in an oxidation step; and L. Removing the leach residue from step J.

62. The method of claim 61 wherein said removed leach residue is removed as a reject tails stream.

63. The method of claim 61 wherein said removed leach residue is removed for further processing.

64. A method of producing copper sulfide from an aqueous solution comprising th steps of: A. Contacting an aqueous solution containing chalcopyrite with sulfur dioxide and a solution of copper sulfate to produce a solid suspension o cuprous sulfide and leach residue; B. Thickening the solid suspension to produce a thickened suspension of cuprous sulfide and leach residue and a clear solution containing sulfuri acid; C. Recycling a portion of the thickened suspension of cuprous sulfide and leach solution to step A;

D. Removing the remaining thickened suspension of cuprous sulfide and leach residue from step C for further processing; E. Contacting the thickened suspension of cuprous sulfide and leach residue with oxygen to produce copper sulfate in solution;

F. Removing residual solids from the copper sulfate solution from step E for further processing;

G. Recycling a portion of the copper sulfate solution from step F to step A; H. Removing the remaining portion of copper sulfate solution from step F for further processing by solvent extraction/electrowinning to produce cathode copper and a copper-containing raffinate; and I. Removing a portion of the raffinate from step H for further processing; J. Processing the remainder of the raffinate from step H by routing to waste water treatment;

K. Returning a part of said portion of said raffinate from step I to step E; L. Routing the remainder of said portion of raffinate from step K to a source of oxidic copper; M. Removing the clear solution containing sulfuric acid from step B and routing at least a portion thereof to said source of oxidic copper;

N. Producing from the source of oxidic copper, raffinate and clear solution a pregnant solution containing copper sulfate; and O. Routing the pregnant solution to the solution extraction/electrowinning process of step H .

65. The method of claim 65 wherein said aqueous solution containing chalcopyritic- bearing solids further contains mixed oxidic materials.