Patent Grant
9725784
The present invention relates to improved methods for production of copper from copper sulfide concentrates produced as part of a mineral ore refining.
De Re Metallica by Georgius Agricola, published in 1556, details the mining, smelting, and refining techniques and technologies of that era. Since then the basic chemical reactions to produce copper have not significantly changed, while the modern smelting process now treats a concentrate rather than as-mined ore of that time. However, technology has markedly advanced through numerous changes and improvements to copper smelting methodology since De Re Metallica's publication. The “Welsh” process, based on a series of sequential reverberatory smelting steps, subsequently dominated copper smelting for over a hundred years. In the 1890s, Nicholls and James developed a process (Great Britain Patent 18,898) based on an alternative final step in the traditional “Welsh” copper smelting process. In this invention part of the high-grade white metal stream was diverted for calcination to produce a copper oxide material for subsequent re-use in the oxidation of the main white metal stream to produce metallic copper. The large, fuel-fired reverberatory furnace was later used for concentrate smelting throughout the first three-quarters of the twentieth century. In more modern times, newer flash and bath smelting processes were developed. The flash smelting concept was described by Bryk et al. in U.S. Pat. No. 2,506,557. Later, Gordon et al described a variant of the flash smelting process in U.S. Pat. No. 2,668,107. An alternative to flash smelting is the bath smelting process such as introduced by McKerrow et al. in U.S. Pat. No. 4,005,856 and also Bailey et al. in U.S. Pat. No. 4,504,309. Still another bath smelting approach, referred to as the Isasmelt process, based on a top lance blowing system with the particular lance system described by Floyd in U.S. Pat. Nos. 3,905,807 and 4,251,271, was developed. The lance system is used in the process operating in Arizona as described by Bhappu et al in: EPD Congress 1994, Edited by G. Warren, The Minerals, Metals and Materials Society, 1993, pages 555 to 570. Each of the contemporary processes described above for the modern era produce a medium to high-grade of copper matte which is typically processed in Peirce-Smith converters to blister copper. Following this, the produced copper is transferred to an anode furnace (European Patent 0648849 B2) for finishing to anode copper for subsequent casting and thence to electrolytic refining. The conventional flash furnace and converter process flow sheet is depicted in
A fossil fuel may be used as a supplementary energy source as required for heating/sustaining typical flash temperatures above 1350° C. A silica flux is added during this step to flux with the iron oxide product shown in Equation (1). The resulting flash furnace slag is sent to a slag treatment facility for copper recovery. The process off-gases are first cleaned and are then treated in a sulfuric acid plant for sulfur recovery.
The remaining molten white metal is transferred to a converter, where it is blasted with oxygen-enriched air to remove remaining sulfides, produce the blister copper, and form an additional slag (Equations 2 and 3).
Cu2S+O2=2Cu+SO2 ΔH°=−59 Wh (2)
FeS+1.5O2=FeO+SO2 ΔH°=−130 Wh (3)
The converter slag is typically higher in copper content, and also requires slag treatment. The flue gases from this step also require processing in the sulfuric acid plant. The copper melt is sent to anode casting (often proceeded by an anode furnace to further purify the copper metal) and then on to electrolysis.
In total, this flash process has gained wide-spread acceptance in the copper industry. Its advantages over older reverberatory molten bath smelting are manifold: utilization of the heat released during oxidation of sulfides with oxygen, high furnace throughput, high copper recovery into matte, and higher SO2 content in the off gas relative to the molten bath process. However, and as previously mentioned, significant control must be maintained throughout the process and significant opportunities for improvement exist. Principally, the composition of the feed materials must be well specified, an understanding of the absolute and relative particle sizes is required, moisture and sulfide contents of the concentrates and fluxes must be quantitatively known, and furnace dimensions and temperatures are critical. Precise control over the feed ratios and rate of oxygen injection must be maintained. Similarly, the amount of siliceous flux that must be added is wholly dependent on the sulfide concentrate and the amount of iron that must be oxidized; high copper losses into the slag are still observed and this requires a separate treatment step. The energy demands of the flash process require preheating of the furnace to circa 900-1100° C. to initiate the exothermic reactions involved when oxygen enrichment is not used. This high temperature conversion leads to NO, formation. Oxygen-enriched air is normally used, in which case preheating the air is not common.
Several variations on flash smelting technology have been developed since the Gordon et al. first work. U.S. Pat. Nos. 5,662,730; 3,790,366; 3,948,639; 3,892,560; 4,615,729; 4,470,845; 3,674,463; 5,607,495; 4,521,245; and US Published Patent Application 2005/0199095 demonstrate oxygen enrichment of air, various techniques for copper recovery from slags as well as partial or dead roasting of the sulfide concentrate prior to flash smelting.
Work performed in the 1890s by Thomas Davies Nicholls, et al. (Great Britain patent 18,898) details the use of copper oxides in roasting copper mattes to copper metal. During this time period, pneumatic copper converting was just in its infancy, hence this method was considered an improvement over the established contemporary roasting process. Copper (I) sulfide, previously smelted into matte (76-78% copper), is crushed and melted in a reverberatory furnace common at that time with calcined copper. The produced copper was then poled to produce a final copper. In this process, it was difficult to produce CuO during the calcination of Cu metal, so Cu2O was used. Production of copper anodes from copper sulfide sources without producing an intermediate copper matte phase has been performed and summarized in the literature1,2. In such operations, the copper sulfide concentrate is first dead roasted at elevated temperatures (900° C.) in an excess of oxygen to produce a copper calcine with sulfur levels around 2% (generally 1-1.5% sulfur). The calcine is then transferred to an electric furnace (e.g. the Brixlegg Process)3,4, a segregation furnaces5,6, a rotary furnace7, or a shaft furnace8,9 where it is further converted to produce blister copper, slag and SO2 off gases. 1Opie W R, (1981) Pyrometallurgical processes that produce blister grade copper without matte smelting. IMM, 137-140.2(1980) Dead Roast-Shaft Furnace copper smelting, World Mining, Vol 33, Issue 12, 40-41.3Kettner P, Maelzer C A, and Schwartz W H, (1972) The Brixlegg Electro-Smelting Process Applied to Copper Concentrates, AIME Annual Meeting, San Francisco.4Paulson D L, Worthington R B, and Hunter W L, (1976) Production of Blister Copper by Electric Furnace Smelting of Dead-Burned Copper Sulfide Concentrates, U.S. Bureau of Mines, RI-8131.5Opie W R, and Coffin L D, (1974) Roasting of Copper Sulfide Concentrates Combined with Solid State Segregation Reduction to Recover Copper, U.S. Pat. No. 3,799,764.6Pinkney E T, and Plint N, (1968) Treatment of Refractory Copper Ores by the Segregation Process, Transactions of AIME, Vol 241, 373-415.7Rajcevic H P, Opie W R, and Cusanelli D C (1978) Production of Blister Copper in a Rotary Furnace from Calcined Copper-Iron Concentrates, U.S. Pat. No. 4,072,507.8Rajcevic H P, Opie W R, and Cusanelli D C (1977) Production of Blister Copper Directly from Dead Roasted-Copper-Iron Concentrates Using a Shallow Bed Reactor, U.S. Pat. No. 4,006,010.9Opie W R, Rajcevic H P, Querijero E R, (1979) Dead Roasting and Blast-Furnace Smelting of Chalcopyrite Concentrates, Journal of Metals, Vol 31, Issue 7, 17-22.
It is an object of the present invention to provide a better method to recover copper from copper sulfide concentrates via a process chemistry previously unused by the copper smelting industry. This process is referred to as the “Looping Sulfide Oxidation” (or “LSO”) process.
Many of the opportunities for improvement in flash smelting outlined above stem from the incremental removal of sulfur in two separate processing steps. As a result, the concentrations of the SO2 streams, which while higher than the concentrations in the roaster and reverberatory furnace off gases (ca. 15-20% SO2), the presence of two sulfurous off gas streams requires handling and treatment, and slags with relatively high copper contents are produced in both the flash furnace and the converter. The Looping Sulfide Oxidation process for copper production removes sulfur in a single step while using copper oxides (Cu2O and CuO) as oxidizing agents to either replace or augment oxygen (O2) from natural air without producing a matte phase. Reference herein to copper oxide oxidizing agents include copper carbonates, sulfates and other oxygen containing copper compounds thermodynamically suitable for use in the Looping Sulfide Oxidation process following the guidelines shown in this application.
Looping Sulfide Oxidation features three distinct steps: conversion of the copper sulfide concentrates into copper and copper oxides (wholesale desulfurization), recovery of copper from the slag, and looping oxide regeneration (
In this first step of the Looping Sulfide Oxidation process, the copper concentrate is blended with fluxes and the oxidizing agent, CuO. In alternative embodiments, the CuO may be augmented with oxygen from air in a fashion such that the total stoichiometry of the system is maintained. The reaction that takes place in this furnace is presented below.
In such a reaction scheme, the value of a is allowed to vary such that ratio of CuO to O2 might range from 5:0 to minimal CuO with greater portions of O2 while still satisfying the reaction stoichiometry. While the relative ratio of CuO and O2 is important, the total amount of oxidizer may be equal to or in excess of the amount required to completely oxidize the copper concentrate. Consideration must be made that excess of the oxidizer can influence the copper melt and/or slag compositions. In this sense, CuO functions to oxidize the iron in the concentrate and/or slag in addition to oxidizing (desulfurizing) the copper in the concentrate. A fraction of copper will be present in the slag as Cu2O due to the equilibrium established between the slag and the copper metal phase. As such, the calculated stoichiometry of the oxidizing agents is minimal, and will be exceeded.
One possible embodiment of the furnace is a Vanyukov-type furnace10,11 (exemplars of which appear in U.S. Pat. Nos. 4,252,560 and 4,294,433), i.e. the concentrate and fluxes are added through the slag, which is agitated by the injection of N2, hot combustion products, and/or air through tuyères; additionally, the energy is supplied via electrodes submerged in the slag. Due to the high energy demand of the endothermic reaction that takes place, additional heat must be provided to the first furnace. This heat will be supplied either solely through the electrical heating of the furnace or through electrical heating augmented by combustion of fuels, whose heat will be transmitted to the furnace through the hot gases in the tuyères and whose chemically inert combustion product gases will be injected into the molten slag to facilitate mixing. Another embodiment may use a top-blown lance in the slag in an Isasmelt-type furnace; this embodiment may also include electrode heating. 10Bystov, V P, Fyodorov, A N, Komkov A A, and Sorokin M L (1992) The use of the Vanyukov process for the smelting of various charges, in Extractive Metallurgy of Gold and Base Metals, Australasian Institute of Mines and Metallurgy, Pardville, Vic., 477-482.11Bystrov, V P, Komkov A A, and Smimov L A (1995) Optimizing the Vanyukov process and furnace for treatment of complex copper charges, in Copper 95-Cobre 95, Vol. IV—Pyrometallurgy of Copper, ed. Chen W J, Diaz C, Luraschi A, and Mackey P J, The Metallurgical Society of CIM, Montreal, Canada, 167-178.
The process chemistry that takes place in the first furnace is of critical importance. Most notably, metal not matte is formed during this step. This marks a significant differentiation and improvement over the present state of the art. The molten copper metal produced in the furnace is very low in iron and is sent directly to the anode furnace. The oxidized iron slag contains copper that must be recovered during slag treatment. As previously mentioned, the complete desulfurization of the concentrate is accomplished in this single step. This allows for significant energy capture during sulfuric acid production in an acid plant. Additionally, because no sulfurous/sulfuric gases will be produced in the downstream processing, aggressive energy capture can be performed on the off gases without fear of acid condensation. The major differences between this invention and the closest prior art (Nicholls et al.) are:
During slag treatment, the goal is to recover as much copper as possible from the slag phase so that it can be returned to the processing loop for copper anode production. In general, the slag from the first furnace will contain ca. 10-15% copper in the slag as Cu2O. The slag, which is still molten, is treated with either carbon (from coal or natural gas) to reduce the copper oxides to copper metal (and the trivalent iron to divalent iron), or oxidized with sulfur (e.g., as iron pyrite), to produce copper matte. With carbon reduction the copper from the slag treatment furnace can be mixed with the copper rich material from the smelting furnace; with sulfidation, the matte will be returned to the smelting furnace to be reprocessed.
Slag treatment must reduce the copper content in the waste slag to levels below ca. 0.4 weight percent. The copper solubility in the slag is a function of many variables; one of critical importance is the Fe(III):Fe(II) ratio. In this process, the copper solubility in the slag is reduced (and thereby the copper recovery is increased) by significantly reducing the Fe(III) content in the slag. Additionally, when the product from slag treatment is copper metal, the iron content must be sufficiently low enough for an anode furnace. In the process presented here, the copper metal from the slag treatment step is blended with the copper metal from the first furnace to produce a copper-rich stream to be processed in the anode furnace. If sulfidation is performed, the copper matte produced will be processed in the first furnace. This step in the process is carried out in a traditional slag treatment furnace, e.g. an electric furnace.
The anode furnace operates in the same fashion as conventional anode furnaces. The copper melt is first oxidized to oxidize any residual iron to a dry slag; in this step some of the copper metal may be co-oxidized. The slag is tapped off and the remaining copper melt is then deoxidized prior to casting to anodes ready for electrolytic refining.
The fraction of copper that is sent to electrolysis is determined by the stoichiometry of the reaction in the smelting furnace (i.e. the amount of copper in the concentrate is equal to the amount of copper in the anodes for electrolysis). The necessary amount of copper to produce the requisite copper oxide for oxidation of the copper concentrate is sent to the reoxidation furnace. In this furnace, the copper melt is atomized and oxidized to CuO with air. This highly exothermic reaction can be harnessed for energy capture. The molten copper is oxidized at high temperatures in a downer or vertical furnace (ca. 1500° C.), and cooled below freezing to ca. 800° C. The powdered CuO is then looped back to the smelting furnace to complete the reaction cycle.
2Cu+O2→2CuO ΔH1500° C.=−59.5 Wh (5)
This invention provides an improvement over the closest prior art wherein Cu2O was produced (Nicholls et al.) in which copper matte is oxidized to produce copper oxide. In this work, copper is reoxidized after atomization to promote rapid and complete oxidation.
Alternatively, other sources of copper oxides can be used as oxygen carriers during Looping Sulfide Oxidation. For example, CuO is used in the industry as pigments in ceramic materials, battery materials and catalysts. These materials can be fed to the smelting furnace to augment the copper oxides that are produced in the reoxidation furnace. Similarly, several copper oxide minerals are processed by the copper industry; these minerals can be used as source of copper oxides during Looping Sulfide Oxidation. Thermodynamic calculations, made with FactSage 6.412 thermodynamic software, detailing such operation are disclosed below. 12Bale, C. W., et al., FactSage™ 6.4.1, Thermfact and GTT-Technologies, CRCT, Montreal, Canada (2013).
Copper scrap is also an important copper stream for Looping Sulfide Oxidation. Copper scrap metals and copper alloy scrap can be processed in Looping Sulfide Oxidation via either smelting in the smelting furnace in the presence of copper oxides (potentially augmented with air), or via initial oxidation to copper oxides in the reoxidation furnace. In the former embodiment, the copper scrap is melted in the smelting furnace and converted to copper metal in the same fashion as copper concentrate. Depending on the composition of the scrap, alloyed metals will report to either the slag or the copper phase. The use of this embodiment can gain an increase in the iron content in the molten copper due to the reduction of the iron oxides present in the slag with any reducing metals (e.g. aluminum or silicon) present in the scrap. In the latter embodiment, the copper scrap is processed to enable its rapid atomization and oxidation (in one embodiment, in a plasma furnace) to copper oxides that can be looped to the smelting furnace.
Other objects, features and advantages will be apparent from the following detailed description of preferred embodiments taken in conjunction with the accompanying drawings in which:
In the present analytical example (not based on a physical plant actually constructed) the process is described on a production basis of approximately 1000 kg of anode copper. The process flow (all or parts of which can be continuous, semi-continuous or batch format) is shown in
Electric Furnace
A room temperature copper concentrate comprising 3000 kg CuFeS2, 173.4 kg FeS2, and 294.8 kg gangue (CaO, Al2O3, SiO2), preferably in free flowing powder form, is to be mixed with 7400 kg of CuO at 800° C. in the first smelting furnace (Table 1). Heat and material balances were calculated using HSC 7.1 Chemistry for Windows thermochemical software13. Silica (1000 kg) and lime (500 kg) fluxes are also taken as to be added to the melt. The melt is to be heated to 1300° C. via electrical and/or combustion heating. The reaction produces a metallic copper melt, an oxidized slag, and a rich SO2 gas stream. In this Example, 14% excess CuO is used to produce an optimal copper melt and an optimal slag (
As discussed in the above Summary of the Invention, the copper solubility in the slag is largely dependent on the degree of oxidation of the iron also present in the slag. The fluxes added to the furnace are designed to aid in slag formation and produce a low melting, fluid slag. The slag produced in this Example melts at 110° C. with a viscosity of 2.0 poise (at 1300° C.). The Cu2O content in the slag is 13.2%, and requires treatment to recover as much of this copper as possible (
The SO2 stream produced during the smelting step is sent to an acid plant for sulfuric acid production. The SO2 content of the off gas in this Example is 46%. Significant energy can be captured during sulfuric acid production, and this energy can be used to improve the overall energy balance of the Looping Sulfide Oxidation process.
Slag Treatment
The slag produced in the electric furnace (3.0% Al2O3, 31.3% SiO2, 15.5% CaO, 23.3% FeO, 13.6% Fe2O3, 13.2% Cu2O) is transferred to an electrical furnace at 1300° C. for slag treatment (Table 2). In this Example the 3596.9 kg of slag is treated with 52 kg of carbon to reduce Fe2O3 and Cu2O. By reducing the trivalent iron, the solubility of copper in the slag is dramatically reduced. As a result, a copper melt is formed with 97.7% of the copper recovered (417.1 kg melt, 98.997% Cu, 1.0% Fe) (
The heat required to perform the slag treatment will be provided by electrical heating via the electric furnace. Natural gas for combustion heating can also be provided via tuyères.
Downer Reoxidation Furnace
In a downer furnace, molten copper is atomized and oxidized in situ to fine particulate CuO. Atomizing the molten copper minimizes mass transfer limitations between the molten copper and the oxygen and leads to near 100% conversion to CuO. This highly exothermic reaction provides significant potential for energy capture. It is understood that molten CuO is highly corrosive, so following oxidation cool air is introduced to solidify the CuO. The CuO is thus cooled down to 800° C. before it exits as a fine particulate and is recycled back at temperature to the first furnace. Looping of this material in this system at temperature and at high processing speed enhances the overall energy balance of the process.
The flue gases are sent to an air/air heat exchanger, where the reaction air for the downer furnace and anode furnace are preheated to 400° C. in order to maximize the thermal efficiency. The flue gas is then sent to a boiler where a significant portion of the energy is captured as high pressure steam.
In this Example, 7400 kg of CuO are required in the electric furnace. As such, 5911.1 kg of molten Cu must be oxidized in the downer reoxidation furnace; the remaining 1020.5 kg of Cu can be sent to electrolysis for final purification (Tables 3 and 4). In the downer reoxidation furnace a significant excess of air will be used to ensure complete reoxidation.
Energy is captured during this step by using the flue gases from the reoxidation furnace to (1) preheat the oxidation air and (2) produce high pressure steam in a boiler after preheating.
Energy Balance
The two primary energy producing steps in the Looping Sulfide Oxidation process are the sulfuric acid production in the acid plant and the reoxidation of the Cu to CuO before it is looped back to the electric furnace. The acid plant per se, is outside the scope of this invention; however, as it is known to those skilled in the art, state-of-the-art processes like the Lurec® process have been shown to capture significant portions of the total energy available during sulfuric acid production14. On this basis, we have evaluated the energy balance of the Looping Sulfide Oxidation process relative to conventional copper processing. 14Daum K H, The Lurec® Process—Key to Economic Smelter Acid Plant Operation, in The Southern African Institute of Mining and Metallurgy Sulfur and Sulfuric Acid Conference 2009, 1-22.
During conventional copper processing, the only major energy producing step is the acid production. It is estimated that the theoretical total amount of energy that can be produced during this step is 54.7 Wh per mole of CuFeS2 processed.
2SO2+O2→2SO3 ΔH600° C.=−54.7 Wh (6)
In this analysis, production of sulfuric acid is estimated to result in the production of 2462 kg of high pressure steam (100 bar, 350° C.) per 1000 kg of Cu produced during heat capture in boilers and cooling jackets (Tables 5 and 6).
Therefore, with all other factors being equal, conventional copper processing and Looping Sulfide Oxidation processing would theoretically produce equal amounts of energy during sulfuric acid production. However, as the Lurec® process states, the higher the strength of the SO2 stream, the greater the energy production; therefore, it can be expected that, in practice, the Looping Sulfide Oxidation process would actually produce more energy than the conventional process due to its high strength SO2 stream. However, if equal energy production is assumed in the acid plant, the only major differentiating factor in energy production will be during the reoxidation of the copper to CuO, which the conventional process does not perform. During reoxidation, the amount of high pressure steam (100 bar, 350° C.) that is estimated to be produced is 4049 kg per 1000 kg of Cu produced (Tables 7 and 8).
Taking into consideration the total estimated energy output during Looping Sulfide Oxidation, the amount of energy available for capture during the reoxidation of the molten copper is approximately 1.64 times greater than the amount available for capture during sulfuric acid production alone. This comparison is vital because during conventional processing, significant energy consumptions and productions have been observed at different processing facilities15. Therefore, on the basis of potential energy available for capture, the Looping Sulfide Oxidation process provides significant improvements over the conventional technology; the increased energy production drastically mitigates the net energy consumption during copper processing. 15Coursol P, Mackey P J, and Diaz C M (2010) Energy Consumption in Copper Sulphide Smelting, in Proceedings of Copper 2010, 1-22.
Using the same feed conditions and smelting furnace parameters as those presented in Example 1, the slag produced in the smelting furnace can be treated in the slag treatment furnace by sulfidation. During sulfidation, iron pyrite (FeS2) is added to the molten slag to sulfidize the copper, causing it to separate out of the slag into a copper matte (
Copper sulfide concentrate (CuFeS2) is smelted with CuCO3 to produce copper metal, iron oxide slag, and rich SO2 off gas (
Copper sulfide concentrate (CuFeS2) is smelted with CuSO4 to produce copper metal, iron oxide slag and rich SO2 off gas (
It will now be apparent to those skilled in the art that other embodiments, improvements, details, and uses can be made consistent with the letter and spirit of the foregoing disclosure and within the scope of this patent, which is limited only by the following claims, construed in accordance with the patent law, including the doctrine of equivalents.
This application claims priority from U.S. provisional patent application 61/662,603 and 61/690,210 both filed Jun. 21, 2012. The full content of the said applications and of all other patents, published patent applications and non-patent publications cited herein are incorporated herein by reference as though set out at length herein.
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| Entry |
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| Kettner, P., et al., The “Brixlegg Electro-Smelting Process” Applied to Copper Concentrates, 1972, p. 26, AIME-Tagung San Francisco. |
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| Number | Date | Country | |
|---|---|---|---|
| 20130340568 A1 | Dec 2013 | US |
| Number | Date | Country | |
|---|---|---|---|
| 61662603 | Jun 2012 | US | |
| 61690210 | Jun 2012 | US |
