PRODUCTION OF COPPER VIA LOOPING OXIDATION PROCESS CROSS-REFERENCE TO RELATED APPLICATIONS AND PUBLICATIONS

This application claims priority from U,S. provisional patent application 61/662,603 and 61/690,210 both tiled June 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.

FIELD OF THE INVENTION

The present invention relates to improved methods for production of copper from copper sulfide concentrates produced as part, of a mineral ore refining,

BACKGROUND OF THE INVENTION

De Re Meta!iiea by Georgius Agricoia, 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 Metaiiica'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, NichoHs and James developed a process (Great Britain Patent 1 8,898) 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 materia! for subsequent re-use in the oxidation of the main white metai 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 US Patent

2,506,557. Later, Gordon et ai described a variant of the flash smelting process in US Patent 2,668,107. An alternative to flash smelting is the bath smelting process such as introduced by MeKerrow et ai. in US Patent 4,005,856 and also Bailey et ai. in US patent 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 US patent os. 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 Figure 1. As shown here, copper concentrate is introduced into the flash smelting furnace (as an example of a modern smelting unit) where the copper sulfide concentrate react with oxygen-enriched air to form a medium grad of matte and a slag. The reaction in the flash furnace can he represented by the following equation (Equation 1 ). Some nitrogen will also be present with the oxygen, depending on the degree of oxygen enrichment.

2CuFeS ₂ + " 0 _% - Cu ₂ S - ~FeS + ? FeO + ~ S0 ₂ Δ» ⁰ = -250 Wh (1 )

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 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).

Cu ₂ S + ~ 2 Cu + SOz AH° = -S9 Wh (2)

FeS + 1. 5 0 ₂ = FeO - SO _z ΔΗ ⁰ = -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 easting (often proceeded by an anode furnace to further purify th 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 reverberatorv molten hath smelting are manifold: utilization of the heat released during oxidation of sulfides with oxygen, high furnace throughput, high copper recovery into matte, and higher SC¾ 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 lite feed ratios and rate of oxygen injection must be maintained, Similarly, the amount of siliceous flux that must be added is whol!y 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 fla sh process require preheating of the furnace to circa 900 - 1 100 °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 also required 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. US Patents 5,662,730; 3,790,366; 3,948,639; 3,892,560; 4,615,729; 4,470,845; 3,674,463; 5,607,495; 4,52.1 ,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 1 890s 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, peeumatic 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 (¾() was used. Production of copper anodes from copper sulfide sources without producing an intermediate copper matte phase has been performed and summarized in the literature ^{1 ,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

' Opie WR, (1981 ) Pyrometailurgicai processes that produce blister grade copper without matte smelting M, 137-140.

² ( 1980) Dead Roast-Shaft Furnace copper smelting, World Mining, Vol 33, issue 12, 40-4 i . calcine is then transferred to an electric furnace (e.g., the Brixfegg Process)'' ^"' , a segregation furnace ^'5,0 , a rotary furnace'', or a shaft furnace ^8,9 where it is further converted to produce blister copper, .slag and SC½ off gases.

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,

SUMMARY OF THE INVENTION

Many of the opportunities tor 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 S0 ₂ streams, which while higher than the concentrations in the roaster and reverberatory furnace off gases (c . 15-20% SO ₂ ) 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 (C¾0 and CuO) as oxidizing agents to either replace or augment oxygen (0 ₂ ) 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 thermodynaraically 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 (Figure 2), This process primarily uses CuO as the oxidizing agent instead of 0> in order to eliminate oxygen-enriched air

³ Kettner P. Maelzer CA, and Schwartz WH, ( 3972) The Brixlegg Electro-Smelting Process Applied to

Copper Concentrates, AIME Annual Meeting, San Francisco.

* Paulson DL, Wortbington RB, and Hunter WL, (1976) Production of Blister Copper by Electric Furnace Smelting of Dead-Burned Copper Sulfide Concentrates, U.S. Bureau of Mines, RI-8333.

5 Opie WR. nd Coffin LD, ( 3974) Roasting of Copper Sulfide Concentrates Combined with Solid State Segregation Reduction to Recover Copper, U.S. Patent 3,799,764,

⁰ Pinkney ET, and Flint N, (3 68) Treatment of Refractory Copper Ores by the Segregation Process,

Transactions of AIME, Vol 24 i , 373-415.

' Rajcevic HP, Opie WR, and Cusanelli DC ( 3978) Production of Blister Copper in a Rotary Furnace from Calcined Copper-Iron Concentrates, U.S. Paten; 4,072,507.

³ Rajcevic HP, Opie WR, and Cusanelli- DC (1977) Production of Blister Copper Directly from Dead Roasted- Copper-Iron Concentrates Using a Shallow Bed Reactor, U.S. Patent 4,006,010,

⁵ Opie WR, Rajcevic HP, Querijero ER, (1979) Dead Roasting and B last-Furnace Smelting of Chaicopyrite Concentrates, Journal of Metals, Vol 31, Issue 7, 1 -22, utilization in the sulfur removal step and to generate energy from the reoxidation of copper downstream. Looping Sulfide Oxidation allows for greater energy capture by performing all the desulfurization of concentrates in a single step. Metal refining and slag treatment are handled simultaneously in the second step. Overall copper yield matches well with recovery levels achieved in the conventional flash process.

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 <¾ might range from 5:0 to minimal CuO with greater portions of 0 ₂ 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 th slag as Cu ₂ 0 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 electric furnace

(exemplars of which appear in U.S. patents 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 Ν;·, hot combustion products, and/or air through tuyeres; additionally, the energy is supplied via electrodes submerged in the slag. Due to the high energy demand of the endotherrnic 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

" Bystov, VP, Fyodorov, AN, Komkov AA, and Sorokin ML (1992) The use of the Vanyukov process for the smelting of various charges, Sn Extractive Metallurgy of Gold and Base Metals, Australasian Institute of Mines and Metallurgy, Parkvil!e, Vic, 477-482.

' ' Bystrov, VP, Komkov AA, and Smirnov LA (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 WJ, Diaz C, Lurascht A, and Mackey PJ, The Metallurgical Society of C!M, Montreal, Canada, 167- 178. furnace ihrough the hot gases in the tuyeres 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 isamelt-iype furnace; this embodiment may also include electrode heating.

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/sulfurie 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 ah) are:

1 , The raw material is neither blister copper nor matte, but is rather copper sulfide concentrate

2, The raw material and copper oxide are simultaneously fed into the molten slag in the smelting furnace; the slag is agitated via the injection of combustion product gases or chemically inert gases

3, The use of oxygen is controlled and special care is taken to ensure that the total oxygen from air and copper oxide does not exceed 20% excess of the required stoichiometric amounts.

Slag composition in the smelting step can be further optimized by changing the amounts of fluxes (CaO, SiO?, Ah s) added to reduce the viscosity, lower the melting temperature, and increase copper recovery into the copper melt, increased calcium oxide will decrease the copper solubiiity in the slag. The amount of Fe?C>3 (i.e. the amount of Fe ⁱ⁺ ) in the slag must be reduced.

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 (¾0. 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 frivalent 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 sulfidaiion, 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(HI):Fe(H) ratio, in this process, the copper solubility in the slag is reduced (and thereby the copper recovery is increased) by significantly reducing the FeiHD 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 is 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, n 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 belo freezing to ca. 800 °C. The powdered CuO is then looped back to the smelting furnace to complete the reaction cycle.

2 CM + 0 ₂ → 2CUO AH _iS00 ° _c ~ -59. 5 Wh (5)

This invention provides an improvement over the closest prior art wherein Cu ₂ G 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 oxidatipn.

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 6.4 detailing such operation are disclosed below.

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 siag 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:

BRIEF DESCRIPTION OF THE DRAWINGS

Figure 1 (Prior Art) shows in block diagram form a generalized process flow chart for flash smelting conversion;

Figure 2 shows in block diagram form a Looping Sulfide Oxidation Process to produce anode copper:

Figure 2a shows schematicall an electric, arc furnace used in the smelting conversion;

Figures 3-8 are traces of thermodynamic data showing calculations of production conditions (CuO) feed variation on output conditions of the copper melt and siag during the smelting step;

³ Bale, C.W., et al FactSage™ 6.4,1, Thermfaci and GTT-Technologies, CRCT, Montreal, Canada (2013). Figures 9-15 show traces of thermody namic data detai ling the slag treatment and output of the slag treatment furnace, the treated slag and the copper melt or copper matte;

Figures 16-19 show traces of thermodynamic data detailing the smelting of CuFeSa with CuC(>3; and

Figures 20-22 show traces of thermodynamic data detailing the smelting of CuFeS ₂ with CuSCv

DETAILED DESCRIPTION OF PREFERRED EMBODIMENTS

Example I

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 he continuous, semi-continuous or batch format) is shown in Figure 2 and the preferred basic configuration of the electric furnace (an arc furnace) is shown in Figure 2a including tuyeres for gas injection into a molten slag formed in the furnace.

Electric Furnace

A room temperature copper concentrate comprising 3000 kg CuFeSj, 173.4 kg FeS , and 294.8 kg gangue (CaO, AI2Q3, SiC^), 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 software ^{1 "} '. 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 SO? gas stream, In this Example, 14% excess CuO is used to produce an optimal copper melt and an optimal slag (Figures 3, 4, and 5) ^!4 . The copper melt is 98.8% copper with 0.002% Fe, and 0.88% S (Figure 6). The slag includes some copper oxide (as C11 ₂ O), iron oxides and gangue and flux derivatives. All compositions herein are weight percent unless otherwise noted.

As discussed in the abo ve 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 tormation and produce a low melting, fluid slag. The slag produced in this Example melts at 1 100 C with a viscosity of 2,0 poise (at 1300° C), The€u>0 content in the slag is 13.2%, and requires treatment to recover as

¹³ Rome, A., et at., HSC 7.1 1 , Outotec, Fori, Finland (201 1).

¹⁴ Bale, C.W., et al., FactSagc™ 6.4.1, Thermfact and GTT-Technologies, CRCT, Montreal, Canada (2013), much of this copper as possible (Figure 7). Figure 7 demonstrates that during smelting, the copper content in the slag is largely independent of the slag composition and operating temperature. However, as shown in comparing Figures 8 and 9, the dramatically higher C partial pressure above the slag in the electric furnace as compared to the <¾ partial pressure above the slag in the slag treatment furnace leads to different slag chemistries. Most notably, the decreased copper solubility in the slag after slag treatment can be explained by considering the lower oxygen partial pressure present in the treatment furnace. This demonstrates that the copper content in the slag can be controlled by the oxygen partial pressure.

Table J. Electric Furnace Heat & Mate ial Balance

INPUT

Temperature Amount, Amount, Amount, Latent Total H,

© kmoi kg Nm3 H, kWh kVVh

Cu Concentrate 25.000 18.853 3468, 198 0.855 0.00 -2180.49

CuFeS2 25.000 16.348 3000.000 0.714 0,00 -864.48

FeS2 25.000 1.445 173 ,400 0.035 0,00 -71 .56

CaO*A _B 0 ₃ *2Si0 ₂ 25.000 1.060 294.798 0.107 0.00 - 1244.44

Recycled CuO 800.000 93.029 7400.000 1.173 1025.04 -3040.27

CuO 800.000 93.029 7400.000 1 , 173 1025.04 -3040.27

Flux 25.000 25 ,559 1500.000 0,534 0.00 -5783.49

Si02 25.000 16.643 ί 000.000 0.385 0,00 -421 1 .01

CaO 25.000 8.916 500.000 0.150 0.00 -1572.47

Heating 25.000 47,972 1243.806 888.658 0.00 0,00

C 25,000 8.326 100.000 0.044 0.00 0.00

02(g) 25.000 8.326 266.412 186.609 0,00 0.00

N2(g) 25 ,000 31.320 877.394 702.005 0.00 0.00

Energy Required 2493.56

OUTPUT

Copper Melt 1300.000 105.915 6607.987 31 .555 1449,45 1544.56

Cu 1300.000 102.724 6527.713 0,729 1417.28 1417.28

Fe 1300.000 0.003 0.142 0.000 0.03 0.03

S 1 300.000 1.814 58.147 0.028 21.91 21.91

0(g) 1300.000 1.374 23.985 30.799 10.22 105.33

Siag 1300.000 47,844 3596.867 0.940 1233.75 -7504.18

A1203 1300.000 1 .059 108.020 0.027 44.85 -448.28

Si02 1300.000 18.759 1 127.1 14 0,434 461 ,01 -4285.28

CaO 1300.000 9.973 559.294 0.167 178.67 - 1580.27

FeO 1300.000 1 1.67 ] 838.491 0.140 240. 1 -641.78

Fe203 1300.000 3.057 488.235 0.093 154.03 -547.63

Cu20 1300.000 3.325 475.714 0.079 1 54.79 -0.93

Flue Gas 1300.000 73.419 3407.221 1645.574 1 143 ,46 -25 1.06

802(g) 1300.000 33.772 2163.415 756.960 634.42 -2150,05

C02(g) 1300.000 8,326 366.412 186.609 152.67 -757.38

N2(g) 1300.000 31 .320 877.394 702.005 356.37 356.37 Table 2. Slag Treatment Fu nace Heat & Material Balance

INPUT

Slsig Furnace

Temperature Amount, Ai!iO!nst, Amount, Latent Total H

© kmol kg Nm3 H, kWb kWh

Slag 1300.000 47.844 3596.867 0.940 1233.75 7504.18

A1203 1300.000 1 .059 108.020 0.027 44.85 -448,28

Si02 1300.000 18.759 1 127, 1 14 0.434 461 .01 4285.28

CaO 1300.000 9.973 559.294 0,167 178.67 1580.27

FeO 1300.000 1 3 .671 838.491 0, 140 240.41 -641 .78

Fe203 1300.000 3.057 488.235 0.093 154.03 -547.63

Cu20 1300.000 3.325 475.714 0.079 154.79 -0.93

edtictant 25.000 4.329 52.000 0.023 0.00 0,00

C 25.000 4.329 52.000 0.023 0.00 0.00

s 25.000 0.000 0.000 0.000 0.00 0.00

Energy Required 76.99

OUTPUT

Copper

Melt 3300.000 6.573 437.129 0.053 90.67 90.68

Cu 1300.000 6.49S 412.946 0.046 89.66 89.66

Fe 1300.000 0,075 4.379 0,00 ί 1 .01 1 .03

0(g) 1300.000 0.000 0,004 0.006 0.00 0.02

Treated -

Slag 1300.000 47.430 3080.021 0.843 1054.22 7298.04

AO 03 1300.000 1.059 108.019 0.027 44,85 -448,28

Si02 1300.000 18,759 1 127, 126 0.434 461 .01 4285.33

CaO 1300.000 9.973 559,297 0.167 178.68 1580.28

FeO 1300.000 17.416 1251.250 0.209 358.75 -957,71

Fe203 1300.000 0. i47 23.551 0.004 7.43 -26.42

Cu20 1300.000 0.075 10.778 0.002 3.5 ? -0.02

Flue Gas 1300.000 4.330 151.739 97.040 62.79 -219.84

CO(g) 1300.000 2.425 67.932 54.358 27.86 -46,60

C02(g) 1300.000 1.904 83.807 42,682 34.92 -173.23

The S(¾ stream produced during the smelting step is sent to an acid plant for sulfuric acid production. The SO? 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% A1 ₂ 0 ₃ , 31 .3% Si0 _2s 1 5.5% CaO,

23.3% FeO, 13.6% Fe ₂ 0 ₃ , 13.2% Cu ₂ 0) is transferred to an electrical furnace at 1300 °C tor slag treatment (Table 2). In this Example the 3596.9 kg of slag is treated with 52 kg of carbon to reduce Fe ₂ <¼ and€¾€>. 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) (Figures 10 and 1 1). The remaining slag contains only 0,35%€¾0 and is fit for disposal as waste (melting temperature, 1070 °C; viscosity 1.7 poise at 1300 °C) (Figures 12 and 13), The copper melt produced during slag treatment is blended with the copper melt from the electric furnace to produce a copper stream (7025.1 kg, 98.798% Cu, 0.062% Fe, 0.828% S, 0.313% O) for treatment in the anode furnace.

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 tuyeres.

Dowser Reoxkiaiion Furnace

In a downer furnace, molten copper is atomized and oxidized w 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 Hue gas is then sent to a boiler where a significant portion of the energy is captured as high pressure steam. Table 3. Reoxsdaiion Heat & Material Balance

INPUT

Reoxidatiofs

Tempera tare Amount, Amount, Amount, Total H,

© ksnoi kg Nm3 H, kWh kWh

Copper 1 00.000 93.105 59 ί 5.483 0.862 1284,49 3285.3 1

Cu 1300.000 93.029 591 1.598 0.660 1283.51 1283.51

Fe 1300.000 0.066 3.676 0,000 0.89 0,89

0(g) Ϊ 300.000 0.009 0.144 0.202 0.07 0.69

S 1300.000 0.002 0.065 0.000 0.02 0.02

Reaction Air 400.000 223 .497 6390.255 4964.539 690.25 690.25

02(g) 400.000 46.514 1488.402 1042.553 150.30 3 50.10

N2(g) 400.000 174,982 4901.853 3923 .986 540.15 540.15

OUTPUT

Copper Oxides 1243.850 46. 14 6661.065 3 .309 3421.58 -502.52

Cu20 1243.850 21 ,882 313 1 .300 0.522 585,95 -438.94

C 20(i) Ϊ 243.850 24.600 3520.000 0,587 832.15 -57,68

Cu20*Fe203 3243.850 0.033 9.965 0.000 3 ,48 -5.90

Flue Gas 1243.850 198.197 5644.748 4442.302 2161 .41 2161.24

N2(g) 1243.850 174.982 4901.853 3921 ,986 1895.51 1895.5 1

02(g) 1243.850 3 > 2- 742.764 520,270 265.86 265.86

S02(g) 1243.850 0.002 0.131 0.046 0.04 -0.13

Table 4, Queiieh Cooling of eosidafios Products

INPUT

Reoxidation - Qa nc Cooling

Total

Temperature Amount, Amount, Amount, Latest H,

© kmol kg Nm3 H, kWh kWh

Copper Oxides 1243,850 46.514 6661.065 1 , 109 1421.58 -502.52

Cu20 1243.850 21 .882 3131 ,100 0.522 585.95 -438.94

Cu20(l) 1243.850 24.600 3520.000 0.587 832.15 -57.68

Cu20*Fe203 1243.850 0.033 9.965 0.000 3.48 -5 ,90

Flue Gas 1243.850 198,197 5644.748 4442.302 23 63.43 2161.24

N2(g) 1243.850 174.982 4901 .853 3921.986 1895,51 1 895.5 3

02(g) 1243.850 23.212 742.764 520.270 265.86 265.86

S02(g) 1243.850 0.002 0.13 1 0.046 0.04 -0.13

Cooling Air 25.000 528.256 1 5240.366 1 1 40.122 0,00 0.00

N2(g) 25.000 43 7.322 1 1690.618 9353.697 0.00 0.00

02(g) 25.000 1 10.934 3549.748 2486.426 0.00 0.00

OUTPUT

Copper Oxides 800.000 93.029 7405.274 1 .172 1026.09 3046.62

CuO 800.000 92.996 7397.400 1.172 1024.68 3039.20

CuO*Fe203 800.000 0,033 7.874 0,000 1.41 -7.42

Flue Gas 800.000 703.196 20140.91 1 15761.146 4705.52 4705.35

N2(g) 800.000 592.305 1 592.471 13275.683 3927.10 3927.1 0

02(g) 800.000 1 10.889 3548.310 2485.418 778.39 778.39

S02(g) 800.000 0.002 0,13 1 0.046 0.02 -0, 15

In this Example, 7400 kg of CuO are required in the electric furnace. As such,

591 1.1 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 boiier 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 production' ⁵ . On this basis, we have evaluated the energy balance of the Looping Sulfide Oxidation process relative to conventional copper processing.

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 moie of CuFeS ₂ processed. 2S0 ₂ 4- 0 ₂ → 250 ₃ °c ^{~ ~S4} - ^{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).

" Daum 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. Table 5. Acid Plant Bolter Heat & Material Balance

INPUT

Acid Plant

System - Boiler

!

Temperature Amount, ΑϊΤΒΟϋϊίί, Amount, Latent Total

© kmoi kg Nsra3 H, kW H, kWh

Gas from

Smelting 1300.000 73 ,419 3407,221 1645.574 1 143.46 2551.06

S02(g) 1300,000 33.772 2163.415 756.960 634.42 2150.05

C02(g) 1300.000 8.326 366.412 186.609 152.67 -757.38

N2(g) 1300.000 31 ,320 877.394 702.005 356.37 356.37

Gas from Anode

Furnace 1200.000 50.3 73 1437.869 1 124.550 580.51 -561 ,04 2(g) 1200.000 35.269 987,991 790.495 367.09 367.09

S02(g) 1200.000 1 .810 1 15.964 40.575 31.08 -1 18.18

02(g) 1200,000 0.003 0.082 0.057 0,03 0,03 O(g) 1200.000 0.001 0.026 0,019 0.01 0.03

S03(g) 1200.000 0.001 0,052 0.01 0.02 -0.06

H20(g) 1200,000 8.273 149,040 385.428 108.08 -447.65

C02(g) 1200.000 3.849 169.393 86.270 64.33 -356.39

CO(g) 1200,000 0.514 14.407 1 1.528 5.41 -10.38

H2(g) 1200.000 0,453 0.914 10,163 4.46 4.46

Cooling Water 25.000 89.461 161 1.655 1.61 7 0.00 7098.80

H2O(! 00barl) 25.000 89.461 161 1.655 1.617 0.00 7098.80

OUTPUT

Gas to Scrubbing 400,000 123.591 4845,090 2770.123 465,76 4370.3 1

S02(g) 400.000 35.583 2279.379 797.534 171.29 2762.43

C02(g) 400.000 12,175 5.35.805 272.879 55,58 1275.20

N2(g) 400,000 66.589 1865,385 1492.500 205.55 205,55

02(g) 400.000 0.003 0.082 0,057 0.01 0.01

NO(g) 400.000 0.001 0,026 0.019 0.00 0.02

S03(g) 400.000 0.001 0.052 0.015 0,00 -0.07

H20(g) 400.000 8.273 149.040 185.428 30.33 ••525.40

CO(g) 400.000 0.534 14.407 1 1.528 1.62 -14.17

H2(g) 400,000 0.453 0.91 10.163 1 ,37 1.37

High Pressure

Steam 350.000 89.461 161 1.655 2005, 140 485.79 5840.59

H2O( 100barg) 350.000 89.461 161 1.655 2005.140 485.79 5840.59 Table 6, Catalyst Bed Heat & Materia! Balance

INPUT

Catalyst Bed

Temperature Amount, A

© Bono! kg Nm3 H, kWb H, kWh

Post-Scrub Gas -

Stream 80,000 1 14,346 4680.570 2562.913 59,08 4205.41

S02(g) 80.000 35.583 2279,379 797,534 22.29 2.91 1 .43

N2(g) 80.000 66.589 1865.385 1492.500 29.65 29.65

C02(g) 80.000 12.375 535.805 272,879 7.14 1323,64

Reaction Air 25.000 93 ,193 2688.636 2088.781 0,00 0.00

02(g) 2.5.000 19.570 626.230 438.644 0.00 0,00

N2(g) 25.000 73.622 2062.406 1 50.137 0.00 0.00

Cooling Water 25.000 47,217 850.633 0.853 0,00 3746.75

H2O(i 00barl) 25.000 47.217 850,633 0.853 0.00 3746,75

OUTPUT

Catalyzed Gas 225,000 189.748 7369.206 4252.926 373,05 4869,49

S03(g) 225.000 35.583 2848.680 797.534 1 14.14 3797.62

N2(g) 22.5,000 140.21 1 3927.791 3142.636 228.08 228.08

C02(g) 225.000 12.175 535.805 272.879 27,84 1302.94

02(g) 225.000 1.779 56.930 39.877 2.98 2.98

High Pressure -

Steam 350,000 47.217 850.633 1058.314 256.40 3082.67

H2O(l G0barg) 350.000 47.217 850.633 1058.314 256.40 3082.67

Therefore, with all other factors being equal, conveniional copper processing and Looping Sulfide Oxidation processing would theoretically produce equal amounts of energy durrag sulfuric acid production, However, as the Lurec© process states, the higher the _; strength of the Si¾ 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 S<¼ 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 Cn produced (Tables 7 and 8). S 9

Fable 7. Reoxsdatiosi R .eaciion Air Prehs ^a er Heat & Material ί Balance

INPUT

Reoxidation Heat

Recovery - Air

Pi'eheater

Latent Total

Temperature Amount, Amount, Amount, H, ₉

© ksnoi kg Nm3 kWh kWh

New Reaction Air 25.000 266.141 7678.264 5965.184 0.00 0.00

02(g) 25.000 55,890 1788,402 1252.689 0,00 0.00

N2(g) 25.000 210.252 5889,862 4712.495 0.00 0.00

Reoxidation Flue Gases 800.000 703.196 20140.91 S 15761 .146 4705.52 4705,35

N2(g) 800.000 592.305 16592.471 13275.683 3927.10 3927.10

02(g) 800.000 1 10.889 3548.310 2485.418 778,39 778.39

S02(g) 800.000 0.002 0.131 0.046 0.02 -0.1

OUTPUT

New Reaction Air 400,000 266.341 7678.264 5965, 184 829.38 829.38

02(g) 400.000 55.890 1788.402 1252.689 180.35 180.35

N2(g) 400.000 210.25 2 5889.862 4712.495 649.02 649.02

Reoxidation Flue Gases 671 ,626 703,194 20140.780 15761.101 ; 3875.97 3875.97

N2(g) 671.626 592.305 16592,471 13275.683 3235.65 3235.65

02(g) 671 626 1 10,88 9 3548.310 2485.418 640.32 640.32

S02(g) 671 ,626 0.000 0.000 0.000 0.00 0.00 fable 8. Reoxidati on Boiler Heat & : Material B

INPUT

Reoxidation Heat

Recovery - Boiler

Amount, A mount. Amount, Latent H, Total H,

Temperature€> kraoi kg Nm3 kWh kWh

Reoxidation Flue

Gases 671.626 703.1 4 20140.780 15761 .301 3875.97 3875.97

N2(g) 671.626 592.305 36592.471 13275.683 3235.65 3235.65

02(g) 671.626 1 10.889 3548.310 2485.418 640.32 640.32

S02(g) 671.626 0.000 0.000 0.000 0.00 0.00

Cooling Water 25.000 224.748 4048.881 4.061 0.00 -17833.96

H2O(100bari) 25.000 224.748 4048.881 4.061 0.00 -37833.96

Reoxidation Flue

Gases 150.000 703.194 20140.780 15761.101 735.05 715.05 2(g) 150.000 592.305 16592.473 13275.683 600.31 600.33

02(g) 150.000 3 10.889 3548.3 10 2485.4 S 8 1 14.73 1 14.73

S02(g) 150.000 0,000 0.000 0.000 0.00 0.00

High Pressure

Steam 350.000 224.748 4048.883 5037,413 1220,43 -14673.04

H2O(100barg) 350.000 224.748 4048.883 5037.413 3220.43 -14673.04 Taking into consideration the total estimated energy output during Looping Sulfide Oxidation, the amount of energy available for capture during the reoxidatson of the molten copper is approximately 1 ,64 times greater than the amount available for capture during sulfuric acid production alone. This comparison is vita! because during conventional processing, significant energy consumptions and productions have been observed at different processing facilities' ⁶ . 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.

Example 2

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 (FeS ₂ ) is added to the molten slag to sulfidize the copper, causing it to separate out of the slag into a copper matte (Figures 14 and 15). in this scheme, the copper recovery from the slag ranges from 99 to 96% in the temperature range of 1200 - 1400 °C. The slag has a melting temperature of 1 120 °C and a viscosity of 0.709 poise at 1300 °C. The treated slag is fit for disposal as waste. The copper matte, which is now rich in copper sulfide, must be processed in the smelting furnace again before the copper can be sent to the anode furnace as blister copper.

Example 3

Copper sulfide concentrate (CuFeS ₂ ) is smelted with Cu€(¾ to produce copper metal, iron oxide slag, and rich SO _? off gas (Figures 16-1 8). In such a reaction, 3000 kg of CuFeSj (with 173.4 kg of FeS ₂ and 294.8 kg of CaAl ₂ Si ₂ O _s ) is reacted with 1 1500 kg of CuC0 ₃ and 1000 kg of Si0 ₂ and 500 kg of CaO between 1200 °C and 1400 °C. The products of this reaction will include an off gas that is comprised mainly of CO _? and S0 ₂ (Figure 19). At 1300 ^C C, 6654 kg of molten Cu will be produced containing 0.30% S, 0.21 % O and 0.0028% Fe, The 3490 kg of slag produced contains 10,7% Cu ₂ 0.

Example 4

Copper sulfide concentrate (CuFeS ₂ ) is smelted with CuSC¼ to produce copper metal, iron oxide slag and rich SC½ off gas (Figure 20-22), In such a reaction, 3000 kg CuFeSa (with 173.4 kg FeS ₂ , 294.8 kg CaAl ₂ Si ₂ 0s) is reacted with 7423 kg CuS0 and

⁵ * Coursol P, Mackey PJ. and Diaz CM (2010) Energy Consumption in Copper Sulphide Smelting, in

Proceedings of Copper 2010, 1-22, 1000 kg Si0 ₂ and 500 kg CaO between 1200 °C and 1400 °C. The products of this reaction will include an off gas that is comprised of S0 ₂ that is diluted with any combustion gases or inert gases. At 1300 °C, 3658 kg of molten copper will be produced containing 1 , i% S, 0.33% O and 0,0025% Fe. The 3559 kg of slag produced contains 12.3% Cu ₂ 0.

it will now be apparent to those skilled in the art thai 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,

What is claimed is: