c© 2005 Wiley-VCH Verlag GmbH & Co. KGaA, Weinheim10.1002/14356007.a07 471

Copper 1

Copper

Adalbert Lossin, Norddeutsche Affinerie Aktiengesellschaft, Hamburg, Federal Republic of Germany

1. Introduction . . . . . . . . . . . . . . . 22. Physical Properties . . . . . . . . . . . 43. Chemical Properties . . . . . . . . . . 64. Occurrence . . . . . . . . . . . . . . . . 84.1. Copper Minerals . . . . . . . . . . . . 84.2. Origin of Copper Ores . . . . . . . . . 94.3. Copper Ore Deposits . . . . . . . . . . 104.4. Copper Resources . . . . . . . . . . . . 104.5. Mining . . . . . . . . . . . . . . . . . . . 105. Production . . . . . . . . . . . . . . . . 115.1. Beneficiation . . . . . . . . . . . . . . . 125.2. Roasting . . . . . . . . . . . . . . . . . . 155.3. Pyrometallurgical Principles . . . . . 175.3.1. Behavior of the Components . . . . . . 175.3.2. Matte . . . . . . . . . . . . . . . . . . . . 175.3.3. Slags . . . . . . . . . . . . . . . . . . . . . 175.3.4. Oxidizing Smelting Processes . . . . . 195.3.5. Proposals . . . . . . . . . . . . . . . . . . 195.4. Traditional Bath Smelting . . . . . . 215.4.1. Blast Furnace Smelting . . . . . . . . . 215.4.2. Reverberatory Furnace Smelting . . . 225.4.3. Electric Furnace Smelting . . . . . . . 225.4.4. Isasmelt Furnace . . . . . . . . . . . . . 245.4.5. Noranda Process . . . . . . . . . . . . . 245.4.6. CMT/Teniente Process . . . . . . . . . 255.4.7. Vanyukov Process . . . . . . . . . . . . 255.4.8. Baiyin Process . . . . . . . . . . . . . . 265.5. Autogenous Smelting . . . . . . . . . . 265.5.1. Outokumpu Flash Smelting . . . . . . 275.5.2. Inco Flash Smelting . . . . . . . . . . . 295.5.3. KIVCET Cyclone Smelting . . . . . . 305.5.4. Contop Matte Smelting . . . . . . . . . 305.5.5. Flame Cyclone Smelting . . . . . . . . 315.6. Discontinuous Matte Conversion . . 32

5.7. Continuous Matte Conversion . . . . 355.7.1. Noranda Process . . . . . . . . . . . . . 355.7.2. Mitsubishi Process . . . . . . . . . . . . 365.7.3. Kennecott/Outokumpu Flash Convert-

ing Process . . . . . . . . . . . . . . . . . 385.8. Direct Blister Smelting . . . . . . . . 385.8.1. Blister Flash Smelting . . . . . . . . . 385.8.2. QS Process . . . . . . . . . . . . . . . . . 395.9. Copper Recycling . . . . . . . . . . . . 395.10. Hydrometallurgical Extraction . . . 406. Refining . . . . . . . . . . . . . . . . . . 456.1. Pyrometallurgical Refining . . . . . . 456.1.1. Discontinuous Fire Refining . . . . . . 456.1.2. Continuous Fire Refining . . . . . . . . 466.1.3. Casting of Anodes . . . . . . . . . . . . 466.2. Electrolytic Refining . . . . . . . . . . 476.2.1. Principles . . . . . . . . . . . . . . . . . . 476.2.2. Practice of Electrorefining . . . . . . . 496.3. Melting and Casting . . . . . . . . . . 506.3.1. Remelting of Cathodes . . . . . . . . . 516.3.2. Discontinuous Casting . . . . . . . . . . 516.3.3. Continuous Casting . . . . . . . . . . . 516.3.4. Continuous Rod Casting and Rolling 516.4. Copper Powder . . . . . . . . . . . . . 526.5. Copper Grades and Standardization 536.6. Quality Control and Analysis . . . . 547. Processing and Uses . . . . . . . . . . 557.1. Working Processes . . . . . . . . . . . 557.2. Other Fabricating Methods . . . . . 567.3. Uses . . . . . . . . . . . . . . . . . . . . . 578. Economic Aspects . . . . . . . . . . . . 589. Environmental Protection . . . . . . 6010. Toxicology . . . . . . . . . . . . . . . . . 6111. References . . . . . . . . . . . . . . . . . 62

1. Introduction

Copper [7440-50-8], the red metal, apart fromgold the only metallic element with a color dif-ferent fromagray tone, has beenknown since theearly days of the human race. It has always beenoneof the significantmaterials, and today it is themost frequently used heavy nonferrous metal.The utility of pure copper is based on its physi-cal and chemical properties, above all, its electri-cal and thermal conductivity (exceeded only by

silver), its outstanding ductility and thus excel-lent workability, and its corrosion resistance (achemical behaviormaking it a half-noblemetal).

Its common alloys, particularly brass andbronze, are of great practical importance(→Copper Alloys). Copper compounds andores are distinguished by bright coloration, espe-cially reds, greens, and blues (→Copper Com-pounds). Copper in soil is an essential trace ele-ment for most creatures, including humans.

2 Copper

Etymology. According to mythology, thegoddess Venus (or Aphrodite) was born onthe Mediterranean island of Cyprus, formerlyKypros (Greek), where copper was exploitedmillennia before Christ. Therefore, in earlytimes the Romans named it cyprium, later calledcuprum. This name is the origin of copper and ofthe corresponding words in most Romance andGermanic languages, e.g., cobre (Spanish andPortuguese), cuivre (French), Kupfer (German),koper (Dutch), and koppar (Swedish).

History [21–24]. The first metals found byNeolithic man were gold and copper, later sil-ver and meteoric iron. The earliest findings ofcopper are presumed to be nearly nine millen-nia old and came from the region near Konya insouthern Anatolia (Turkey). Until recently thesix-millennia-old copper implements from Iran(Tepe Sialk) were presumed to be the oldest. IntheOldWorld, copper has beenworked and usedsince approximately

7000 b.c. Anatolia4000 b.c. Egypt, Mesopotamia, Palestine, Iran,

and Turkestan3000 b.c. Aegean, India2600 b.c. Cyprus2500 b.c. Iberia, Transcaucasia, and China2200 b.c. Central Europe2000 b.c. British Isles1500 b.c. Scandinavia

Empirical experience over millennia has ledto an astonishing knowledge of copper metallur-gical operations:

1) Native copper was hardened by hammering(cold working) and softened by moderateheating (annealing).

2) Heating to higher temperatures (charcoal andbellows) produced molten copper and madepossible the founding into forms of stone,clay, and later metal.

3) Similar treatment of the conspicuously col-ored oxidized copper ores formed coppermetal.

4) The same treatment of sulfide copper ores(chalcopyrite), however, did not result in cop-per metal, but in copper matte (a sulfidic in-termediate). Not before 2000 b.c. did peoplesucceed in converting the matte into copperby repeated roasting and smelting.

5) In early times, bronze (copper – tin alloy)waswon from complex ores, the Bronze Age be-

ginning ca. 2800 b.c. At first, copper oreswere smelted with tin ores; later, bronze wasproduced frommetallic copper and tin. Brass(copper – zinc alloy)was knownca. 1000 b.c.and became widely used in the era of the Ro-man Empire.

In Roman times, most copper ore was minedin Spain (Rio Tinto) and Cyprus. With the fall ofthe Roman Empire, mining in Europe came to avirtual halt. In Germany (Saxony), mining activ-ities were not resumed until 920a.d. During theMiddle Ages, mining and winning of metals ex-panded from Germany over the rest of Europe.In the middle of the 16th century, the currentknowledge of metals was compiled in a detailedpublication [23] by Georgius Agricola, DeRe Metallica (1556).

Independent of the Old World, the Indians ofNorth America had formed utensils by workingnative copper long before the time of Christ, al-though the skills of smelting and casting wereunknown to them. On the other hand, the skill ofcopper casting was known in Peru ca. 500 a.d.,and in the 15th century the Incas knew how towin the metal from sulfide ores.

Around 1500, Germany was the world leaderin copper production, and the Fugger familydominated world copper trade. By 1800, Eng-land had gained first place, processing ores fromher own sources and foreign pits intometal. Near1850,Chile became themost important producerof copper ores, and toward the end of the lastcentury, the United States had taken the worldlead in mining copper ores and in production ofrefined copper.

Technical development in the copper indus-try has made enormous progress in the last 120years. The blast furnace, based on the oldestprinciple of copper production, was continuallydeveloped into more efficient units. Neverthe-less, after World War I, it was increasingly re-placed by the reverberatory furnace, first con-structed in the United States. Since the end ofWorld War II, this furnace has been supersededslowly by the flash smelting furnace inventedin Finland. Recently, several even more mod-ern methods, especially fromCanada and Japan,have begun to compete with the older processes.

An important development in producingcrude metal was the application of the Besse-mer converter concept to copper metallurgy by

Copper 3

Manhes and David (France, 1880): this prin-ciple is still the most widely used method forcopper converting in the world.

Over time the requirements for copper pu-rity have become increasingly stringent. The in-vention and development of electrolysis by J. B.Elkington (England, 1865) and E.Wohlwill(Germany, 1876) made refining of high-puritycopper possible.

In addition, the quantity of copper producedhas increased immensely (Table 1). Since 1800,ca. 375× 106 t of primary copper has beenmined in theworld, but of this only ca. 10× 106 twas mined between 1800 and 1900.

Table 1.World mine production of copper (approximate, from sev-eral sources)

Year Production, 103 t Year Production, 103 t

1700 9 1970 64001800 17 1975 73001850 57 1980 79001900 450 1985 83001950 2500 1990 92251955 3100 1995 10 0501960 4200 1997 11 5251965 5000

2. Physical Properties

Most properties of copper metal depend on thedegree of purity and on the source of the metal.Variations in properties are caused by

1) Grade of copper, i.e., the oxygen con-tent: tough-pitch copper, deoxidized copper,oxygen-free copper

2) Content of native impurities (e.g., arsenic,bismuth) or remnants of additives (e.g., phos-phorus), which form solid solutions or sepa-rate phases at the grain boundaries

3) Thermal and mechanical pretreatment of themetal, which lead to states such as cast cop-per, hot-rolled copper, cold-worked (hard)copper, annealed (soft) copper, and sinteredcopper

These property differences are caused by thedefects in the crystal lattice. Two groups of prop-erties are to be distinguished:

1) Low dependence on crystal lattice defects,e.g., caloric and thermodynamic properties,magnetic behavior, and nuclear characteris-tics

2) High dependence on defects, e.g., electricaland thermal conductivity, plastic behavior,kinetic phenomena, and resistance to corro-sion

The variations in properties are caused ei-ther by physical lattice imperfections (disloca-tions, lattice voids, and interstitial atoms) or bychemical imperfections (substitutional solid so-lutions).

Atomic and Nuclear Properties. Theatomic number of copper is 29, and the atomicmass Ar is 63.546± 0.003 (IUPAC, 1983).Copper consists of two natural isotopes, 63Cu(68.94 %) and 65Cu (31.06 %). There are alsonine synthetic radioactive isotopes with atomicmasses between 58 and 68, of which 67Cu hasthe longest half-life, ca. 58.5 h.

Crystal Structure. At moderate pressures,copper crystallizes from low temperatures up toits melting point in a cubic closest-packed (ccp)lattice, type A 1 (also F1 or Cu) with the co-ordination number 12. X-ray structure analysisyields the following dimensions (at 20 C):

Lattice constant 0.36152 nmMinimum interatomic distance 0.2551 nmAtomic radius 0.1276 nmAtomic volume 7.114 cm3/mol

There is also a high-pressure modification,which forms at ca. 400MPa and 100 C.

Density. The theoretical density at 20 C,computed from lattice constant and atomic massis 8.93 g/cm3. The international standard wasfixed at 8.89 g/cm3 in 1913 by the IEC (Interna-tional Electrotechnical Commission). The max-imum value for 99.999 % copper reaches nearly8.96 g/cm3.

The density of commercial copper dependson its composition, especially the oxygen con-tent, its mechanical and thermal pretreatment,and the temperature. At 20 C, a wide range ofvalues are found:

Cold-worked and annealed copper 8.89 – 8.93 g/cm3

Cast tough-pitch electrolytic copper 8.30 – 8.70 g/cm3

Cast oxygen-free electrolytic copper 8.85 – 8.93 g/cm3

4 Copper

The values for cold-worked copper are higherthan those of castings because the castings havepores and gas cavities.

The density of copper is nearly a linear func-tion of temperature, with a discontinuity at themelting point:

Temperature, C Density, g/cm3

solid copper20 8.93600 8.68900 8.47

1 083 8.32liquid copper1 083 7.991 200 7.81

The solidification shrinkage is 4 %; the spe-cific volume at 20 C is 0.112 cm3/g.

Mechanical Properties. Important mechan-ical values are given in Table 2. High-purity cop-per is an extremely ductile metal. Cold workingincreases the hardness and tensile strength (hardor hard-worked copper); subsequent annealingeliminates the hardening and strengthening sothat the original soft state canbe reproduced (softcopper). The working processes are based onthis behavior (Section 7.1). Impurities that formsolid solutions of the substitutional type likewiseincrease hardness and tensile strength.

Table 2.Mechanical properties of copper at room temperature

Property Unit Anealed Cold-worked(soft) copper (hard) copper

Elastic modulus GPa 100 – 120 120 – 130Shearing modulus GPa 40 – 45 45 – 50Poisson’s ratio 0.35Tensile strength MPa 200 – 250 300 – 360Yield strength MPa 40 – 120 250 – 320Elongation % 30 – 40 3 – 5Brinell hardness (HB) 40 – 50 80 – 110Vickers hardness (HV) 45 – 55 90 – 120Scratch hardness ≈3

Pure copper has outstanding hot workabilitywithout hot brittleness, but the high-temperaturestrength is low. Detrimental impurities, thosethat decrease the strength at high temperatures,are principally lead, bismuth, antimony, sele-nium, tellurium, and sulfur. The concentrationof oxides of such elements at the grain bound-aries during heating causes the embrittlement.However, such an effect can be desirable whenfree cutting is required.At subzero temperatures,

copper is a high-strength material without coldbrittleness.

The changes in typical mechanical propertiessuch as tensile strength, elongation, and hard-ness by heat treatment result from recrystalliza-tion [25]. The dependence of recrystallizationtemperature and grain size on the duration ofheating, the amount of previous cold deforma-tion, and the degree of purity of copper can bedetermined fromdiagrams. The recrystallizationtemperature is ca. 140 C for high-purity copperand is 200 – 300 C for common types of copper.A low recrystallization temperature is usuallyadvantageous, but higher values are required tomaintain strength and hardness if the metal isheated during use.

Thermal Properties. Important thermal val-ues are compiled in Table 3. The thermal con-ductivity of copper is the highest of all metalsexcept silver.

Table 3. Thermal properties of copper

Property Unit Value

Melting point K 1356 (1083 C)Boiling point K 2868 (2595 C)Heat of fusion J/g 210Heat of vaporization J/g 4810Vapor pressure (at mp) Pa 0.073Specific heat capacityat 293K (20 C)and 100 kPa (1 bar) J g−1 K−1 0.385

at 1230K (957 C)and 100 kPa 0.494

Average specific heat273 – 573K (0 – 300 C)at 100 kPa (1 bar) J g−1 K−1 0.411

273 – 1273K (0 – 1000 C)at 100 kPa 0.437

Coefficient of linear thermalexpansion273 – 373K (0 – 100 C) K−1 16.9× 10−6

273 – 673K (0 – 400 C) 17.9× 10−6

between 273 and 1173K(0 – 900 C) 19.8× 10−6

Thermal conductivityat 293K (20 C) Wm−1 K−1 394

Electrical Properties. In practice, the mostimportant property of copper is its high elec-trical conductivity; among all metals only sil-ver is a better conductor. Both electrical con-ductivity and thermal conductivity are con-nected with theWiedemann – Franz relation andshow strong dependence on temperature (Ta-ble 4). The old American standard, 100 %

Copper 5

IACS (International Annealed Copper Stan-dard), corresponds to 58.0MS/m at 20 C, andit is still widely used in the United States.The corresponding electrical resistivity () is1.7241×10−8 Ω · cm, and the less usual resistiv-ity based on weight (density of 8.89 g/cm3, IEC)is 0.1533Ω gm−1. The corresponding temper-ature coefficients are 0.0068 ×10−8 ΩmK−1

(d/dT ) and 0.00393K−1 (−1d/dT ). Thetheoretical conductivity at 20 C is nearly60.0MS/m or 103.4 % IACS, and today com-mercial oxygen-free copper (e.g., Cu-OF) has aconductivity of 101 % IACS.

Table 4. Temperature dependence of thermal and electrical conduc-tivity of copper

Temperature Thermalconductivity,Wm−1 K

Electricalconductivity,MS/m

K C

17 −256 5 00073 −200 574 460113 −160 450173 −100 435 110273 0 398 60293 20 394 58373 100 385 44473 200 381 34573 300 377 27973 700 338 15

The factors that increase the strength de-crease electrical conductivity: cold working andelements that form solid solutions. Elementsthat form oxidic compounds that separate atgrain boundaries affect electrical properties onlyslightly. Coppermay lose up to ca. 3%of its con-ductivity by cold working; however, subsequentannealing restores the original value. There is asimple rule: the harder the copper, the lower isits conductivity.

Other Properties. High-purity copper isdiamagnetic with a mass susceptibility of− 0.085× 10−6 cm3/g at room temperature.The dependence on temperature is small. How-ever, a very low content of iron can strongly af-fect the magnetic properties of copper.

The lower the frequency of light, the higherthe reflectivity of copper. The color of a clean,solid surface of high-purity copper is typicallysalmon red.

The surface tension of molten copper is11.25× 10−3N/cmat 1150 C, and the dynamicviscosity is 3.5× 10−3 Pa · s at 1100 C.

Detailed physical-property information anddata are to be found in the literature, particularlyas tabular compilations [25–30].

3. Chemical Properties

In the Periodic Table copper is placed in thefirst transition series (period 4). It belongs toGroup 11 and, together with silver and gold,forms the coinage metals. Its electron config-uration is [Ar] 3d10 4s1. Copper compounds areknown in oxidation states ranging from+1 to +4,although the +2 (cupric) and the +1 (cuprous) areby far the most common. In aqueous solutionsor below 800 C, the +2 oxidation state is themost stable.

Copper(I) compounds such as CuCl and CuIare diamagnetic colorless materials, except forthose whose color results from charge-transferbands, for example, Cu2O. Cu+ ions, [Ar] 3d10,are coordinated in a linear (two ligands) or tetra-hedral fashion (four ligands).

Copper(II) compounds such as CuSO4· 5H2O are paramagnetic blue or green sub-stances, the color of which results from strongabsorption bands in the region between 600 and900 nm caused by d – d electron transfer pro-cesses. The Cu2+ ion is a d9 system and gener-ally sixfold coordinated in a distorted octahedralmanner.

Copper(III) compounds are mostly diamag-netic. Cuprates like NaCuO2 can be obtainedby heating the oxides in pure oxygen. In chem-istry only a few Cu3+ complexes are known, butit appears that Cu3+ plays an important role inbiochemistry, especially with deprotonated pep-tides.

Copper(IV) compounds are not well knownexcept for Cs2[CuF6].

Behavior in Air. Copper in dry air at roomtemperature slowly develops a thin protectivefilm of copper(I) oxide [1317-39-1]. On heatingto a high temperature in the presence of oxygen,copper forms first copper(I) oxide, then cop-per(II) oxide [1317-38-0], both of which coverthe metal as a loose scale.

6 Copper

In the atmosphere, the surface of copper oxi-dizes in the course of years to a mixture of greenbasic salts, the patina, which consists chieflyof the basic sulfate, with some basic carbonate.(In a marine atmosphere, there is also some ba-sic chloride.) Such covering layers protect themetal.

Behavior versus Diverse Substances.While many substances scarcely react withcopper under dry conditions, the rate of at-tack increases considerably in the presence ofmoisture. Copper has a high affinity for freehalogens, molten sulfur or hydrogen sulfide.

Standard electron potentials of copper are asfollows [31], [32]:

Potentials in standard (acid) solution:

Cu+ + e− −→ Cu E0 = 0.521V

Cu2+ + 2 e− −→ Cu E0 = 0.153V

Potentials with complexing ligands:

[Cu(NH3)4]2+ + 2 e− −→ Cu+ 4NH3 E0 =− 0.11V

[Cu(CN)2]− + e− −→ Cu+ 2 CN− E0 =− 0.43V

As the standard electron potentials show, coppermetal is stable to nonoxidizing acids like dilutesulfuric or hydrochloric acid, similar to the pre-cious metals. Dissolution of copper is possiblein oxidizing acids such as nitric acid or hot con-centrated sulfuric acid.Also other redox systemssuch as iron(III) or copper(II) chloride solutionsare suitable reagents for leaching copper in prac-tice.

Copper dissolves not only in oxidizing acidsbut also, for example, in ammonia or cyanide so-lutions in the presence of oxygen because stablecomplexes are formed. Also acetic acid togetherwith oxygen or hydrogen peroxide attacks cop-per forming a green pigment called verdigris.

Free Cu+ ions are not stable in aqueous solu-tion although Cu+ (3d10) has a filled d shell.Spontaneous disproportion into Cu2+ and Cutakes place.

2Cu+ −→ Cu2+ +Cu E0 = 0.37V

K = [Cu2+]/[Cu+] = 106

The distorted octahedral coordination of sixwa-ter molecules around the Cu2+ ion (d9) givesan additional stabilization energy (ligand-fieldeffect). In aqueous solutions, Cu+ is only ex-istent in form of very stable complexes like[Cu(CN)2]− or in the presence of an excess ofcopper metal. Also, insoluble Cu+ compoundssuch as cuprous oxide do not disproportionatein water.

By virtue of its large ionic radius and lowelectrical charge, the Cu+ ion is a soft acid.Therefore, the chemistry of copper in the oxida-tion state + 1 is predominated by reactions withsoft bases like iodine (CuI), sulfur (CuSCN),or unsaturated nitrogen ligands. In contrast,the chemistry of Cu2+, which is smaller andmore highly charged, is dominated by hard lig-ands like oxygen ([Cu(H2O)6]2+) or nitrogen([Cu(NH3)4]2+).

Copper is very stable in fresh water andalso in sea water or alkali metal hydroxide so-lutions. Wastewater containing organic sulfurcompounds can be corrosive to copper.

Figure 1.Pourbaix diagram for copper in highly dilute aque-ous solution at normal temperature [35]

Corrosion [33], [34].M. J. N. Pourbaixhas developed potential – pH equilibrium dia-grams for metals in dilute aqueous solutions[35]. Such graphs give a rough indication of thefeasibility of electrochemical reactions. Figure 1shows the behavior of copper at room temper-ature and atmospheric pressure. The Cu –H2Osystem contains three fields of different charac-ter:

Copper 7

1) Corrosion, in which the metal is attacked2) Immunity, in which reaction is thermody-

namically impossible3) Passivity, in which there is no reaction be-

cause of kinetic phenomena

Gases and Copper [36], [37]. An exactknowledge of the behavior of solid and liquidcopper toward gases is important for productionand use of the metal. With the exception of hy-drogen, [1333-74-0], the solubility of gases inmolten copper follows Henry’s law: the solubil-ity is proportional to the partial pressure.

Oxygen [7782-44-7] dissolves inmolten cop-per as copper(I) oxide up to a concentration of12.6wt % Cu2O (corresponding to 1.4wt % O)(also see Fig. 32). Copper(I) oxide in solid cop-per forms a separate solid phase.

Sulfur dioxide [7446-09-5] dissolves inmolten copper and reacts:

6Cu + SO2 Cu2S+ 2Cu2O

Hydrogen is considerably soluble in liquid cop-per, and after solidification some remains dis-solved in the solid metal, although copper doesnot form a hydride. The solubility follows Siev-ert’s law, being proportional to the square rootof the partial pressure because the H2 moleculesdissociate into H atoms on dissolution. Hydro-gen has high diffusibility because of its ex-tremely small atomic volume.

Hydrogen dissolved in oxygen-bearing cop-per reacts with copper(I) oxide at high tempera-tures to form steam:

Cu2O+2H−→←− 2Cu +H2O(g)

Steam is not soluble in copper; therefore, it ei-ther escapes or forms micropores.

Nitrogen, carbonmonoxide, and carbondiox-ide are practically insoluble in liquid or solidcopper. Hydrocarbons generally do not reactwith copper. An exception is acetylene, whichreacts at room temperature to form the highlyexplosive copper acetylides Cu2C2 and CuC2;therefore, acetylene gas cylinders must not beequipped with copper fittings.

4. Occurrence

In the upper part of the earth’s crust (16 kmdeep), the average copper content is ca. 50 ppm.

It is thus about half as abundant as chromium,about twice as abundant as cobalt, and 26th inorder of abundance of the elements in the acces-sible sphere of the earth. Table 5 shows averagecopper contents in natural materials.

Table 5. Typical copper contents of natural materials

Mineral Content, ppm

Basalt 85Diorite 30Granite 10Sandstone 1Copper ores (poor) 5 000Copper ores (rich) 50 000Native copper 950 000Seawater 0.003Deep-sea clays 200Manganese nodules 10 000Marine ore sludges 10 000Earth’s crust (average) 50Meteorites (average) 180

4.1. Copper Minerals

More than 200 minerals contain copper in defin-able amounts, but only about 20 are of impor-tance as copper ores (Table 6) or as semipreciousstones (turquoise and malachite). Copper is atypical chalcophilic element; therefore, its prin-cipal minerals are sulfides, mostly chalcopyrite,bornite, and chalcocite, often accompanied bypyrite, galena, or sphalerite.

Secondary minerals are formed in sulfide orebodies near the earth’s surface in two stages.In the oxidation zone, oxygen-containing wa-ter forms copper oxides, basic salts (basic car-bonates and basic sulfates), and silicates. In thedeeper cementation zone, copper-bearing solu-tions from these salts are transformed into sec-ondary copper sulfides (chalcocite and covel-lite) and even native copper of often high purity,e.g., in the Michigan copper district (KeweenawPeninsula).

Other metallic elements frequently found incopper ores are iron, lead, zinc, antimony, and ar-senic; less common are selenium, tellurium, bis-muth, silver, and gold. Substantial enrichmentssometimes occur in complex ores. For example,ores from Sudbury, Ontario, in Canada containnickel and copper in nearly the same concentra-tions, as well as considerable amounts of plat-inum metals. The copper ores from Zaire and

8 Copper

Table 6. The most important copper minerals

Mineral Formula Copper, wt % Crystal system Density, g/cm3

Native copper Cu 99.92 cubic 8.9Chalcocite Cu2S 79.9 orthorhombic 5.5 – 5.8Digenite Cu9S5 78.0 cubic 5.6Covellite CuS 66.5 hexagonal 4.7Chalcopyrite CuFeS2 34.6 tetragonal 4.1 – 4.3Bornite Cu5FeS4/Cu3FeS3 55.5 – 69.7 tetragonal 4.9 – 5.3Tennantite Cu12As4S13 42 – 52 cubic 4.4 – 4.8Tetraedrite Cu12Sb4S13 30 – 45 cubic 4.6 – 5.1Enargite Cu3AsS4 48.4 orthorhombic 4.4 – 4.5Bournonite CuPbSbS3 13.0 orthorhombic 5.7 – 5.9Cuprite Cu2O 88.8 cubic 6.15Tenorite CuO 79.9 monoclinic 6.4Malachite CuCO3 · Cu(OH)2 57.5 monoclinic 4.0Azurite 2 CuCO3 · Cu(OH)2 55.3 monoclinic 3.8Chrysocolla CuSiO3 · nH2O 30 – 36 (amorphous) 1.9 – 2.3Dioptase Cu6 [Si6O18 ] · 6H2O 40.3 rhombohedral 3.3Brochantite CuSO4 · 3Cu(OH)2 56.2 monoclinic 4.0Antlerite CuSO4 · 2Cu(OH)2 53.8 orthorhombic 3.9Chalcanthite CuSO4 · 5H2O 25.5 triclinic 2.2 – 2.3Atacamite CuCl2 · 3Cu(OH)2 59.5 orthorhombic 3.75

Zambia are useful sources of cobalt. Many por-phyry copper ores inAmerica contain significantamounts of molybdenum and are the most im-portant single source of rhenium. The extractionof precious metals and other rare elements canbe decisive for the profitability of copper mines,smelters, and refineries.

4.2. Origin of Copper Ores

Ore deposits are classified according to theirmode of formation, but the origin of copper oresis geologically difficult to unravel, and someof the proposed origins are controversial. Theclassification distinguishes twomain groups, themagmatic series and the sedimentary series.

Magmatic ore formation involves magmacrystallization and comprises the followinggroups:

1) Liquid magmatic ore deposits originate bysegregation of the molten mass so that theheavier sulfides (corresponding to matte)separate from the silicates (correspondingto slag) and form intrusive ore bodies. Ex-amples: Sudbury, Ontario; Norilsk, westernSiberia.

2) Pegmatitic – pneumatolytic ore deposits de-velop during the cooling of magma to ca.374 C, the critical temperature of water. Ex-amples: Bisbee, Arizona; Cananea, Mexico.

3) Hydrothermal ore deposits result by furthercooling of the hot, dilute metal-bearing solu-tions from ca. 350 C downward, i.e., belowthe critical temperature of water. Such de-posits contain copper primarily as chalcopy-rite and satisfy ca. 50 % of the demand inthe Western world. There are many exam-ples of different types of hydrothermal de-posits. Examples: Butte, Montana (ganguedeposit); Tsumeb, Namibia (metasomaticdeposit); BinghamCanyon, Utah; Chuquica-mata, Chile; Toquepala, Peru; Bougainville,Solomon Islands (impregnation deposits).Impregnation deposits are also called dis-seminated copper ores or porphyry copperores (or simply porphyries) because of theirfine particle size.

4) Exhalative sedimentary ore deposits origi-nate from submarine volcanic exhalationsand thermal springs that enter into seawa-ter, and constitute a transitional type to sedi-mentary deposits. These ores are third in eco-nomic importance in the Western world. Theactual formation of such sulfidic precipita-tions can be observed, for example, the ma-rine ore slimes in the Red Sea. Examples:Mount Isa, Queensland; Rio Tinto, Spain;Rammelsberg (Harz), Federal Republic ofGermany.

The origin of sedimentary ore occurs in theexogenous cycle of rocks andmay be subdividedinto the following groups:

Copper 9

1) Arid sediments in sandstones and conglom-erates occur widely in the former SovietUnion as widespread continental zones ofweathering with uneven mineralization. Ex-amples: Dsheskasgan, Kazakhstan; Exotica,Chile.

2) Partly metamorphized sedimentary ores inshales, marls, and dolomites form largestrata-bound ore deposits, especially in theAfrican copper belt, and represent the secondmost important source of copper to theWest-ernworld, aswell as supplyingnearly 75%ofits cobalt. Examples: Zaire (oxidation zone,oxidized ores 6 %Cu); Zambia (cementa-tion zone, secondary sulfide ores 4 %Cu).

3) Marine precipitates have formed sedimen-tary ore deposits similar to the present phe-nomenon of sulfide precipitation by sulfurbacteria in the depths of the Black Sea. Ex-amples: Silesia (copper marl), Poland.

4) Deep-sea concretions lie in abundance on thebottom of the oceans, especially the PacificOcean. These so-called manganese nodulescould also be a source of copper in the future.

4.3. Copper Ore Deposits

Geologically, the main regions of copper ore de-posits are found in two formations: the Precam-brian shields and theTertiary foldmountains andarchipelagos. There are major producing coun-tries on every continent [38], [39].

North America: United States (Arizona, Utah,New Mexico, Montana, Nevada, and Michi-gan), Canada (Ontario, Quebec, BritishColumbia, and Manitoba), and Mexico(Sonora)

South America: Chile, Peru, Argentina, andBrazil

Africa: Zaire, Zambia, Zimbabwe, SouthAfrica, and Namibia

Australia and Oceania: Queensland, Papua NewGuinea

Asia: Former Soviet Union (Siberia, Kaza-khstan, and Uzbekistan), Japan, Philippines,Indonesia, India, Iran, and Turkey

Europe: Poland (Silesia), Yugoslavia, Portugal,Bulgaria, Sweden, and Finland

Antarctica may be an important source ofcopper ores in the foreseeable future.

4.4. Copper Resources

World primary copper reserves were estimatedin 1991 at 552× 106 t (Table 7) [41]. Reservesare identified resources and do not includeundiscovered resources. With time the avail-able reserves have increased because of techni-cal progress in processing ores with low coppercontent and the discovery of new ore deposits[38], [39]. About 321× 106 t were classified asminable copper ores under the technical and eco-nomical conditions at that time.

It is believed that a large potential of as-yetuntouched deposits exist. Therefore, the poten-tially usable copper resources are estimated tobe about three times as large as the reserve base.In addition to ores on land, there is an estimatedamount of copper in deep-sea nodules of about0.7× 109 t.

Table 7. Copper ore reserves in 1991 [41]

Country Ore reserves, Percentage of106 t world reserves

United States 90 16.3Chile 120 21.7Peru 31 5.6Zambia 30 5.4Zaire 30 6.3Canada 23 4.2Australia 21 3.8Philippines 16 2.9Indonesia 8 1.4China 8 1.4Poland 15 2.7CIS 54 9.8Other countries 106 19.2

If one assumes that total production of pri-mary copper will remain stable, the identifiedreserves would last until 2040. With an increasein copper production of 2 – 3 %, which is morerealistic, the duration of the known reserves willbe reduced. However, these forecasts are quiteunreliable because the growing use of secondarycopper (recycling materials) and the discoveryof new copper ore deposits are not considered.

4.5. Mining

Exploration, which is the search for ore depositsand their detailed investigation, is required to

10 Copper

ascertain the commercial feasibility of a poten-tial mine. The used geological, geochemical, orgeophysical methods are complicated and ex-pensive. But often legal and political factors aremore decisive for a new mine project than tech-nological or financial aspects. Themethods usedfor exhausting the copper ore are depending toa great extent to the type of deposit. Importantparameters are metal content of the ore and ge-ometry and depth of the ore body. These pa-rameters determine the workingmethod. Under-ground and open-cast mining are the two basictechniques. As a generalization mining can bedivided into the stages drilling, blasting, load-ing, and haulage of the ore.

Technological developments like the LHDtechnique (load, haul, dump) and the use ofconcrete for stowing have made mining morecost efficient. But underground mining is stillmore labor-intensive and expensive than opencast mining. Therefore, copper ores with an av-erage Cu content of 1 % or more or that con-tain other valuable metals in addition to copper(e.g., precious metals, nickel, cobalt) are minedunderground, while ores with 0.5 % Cu, whichrepresents a 100-fold enrichment of the averagecopper content in the earth’s crust, are minedby open-cast methods. The porphyry copper de-posits which are located near the surface canonly be exploited by open-cast technique. Theyhave become more important during the lastdecades. The first open pit was started at Bing-ham Canyon early this century. Today the ter-raced copper open pits are the largest ore minesin the world. They often cover more than onesquare kilometer and have working depths ofseveral hundred meters. 100 000 t of crude oreare extracted per day. Today more than 50 % ofthe primary copper comes from open pits. Otherless common methods for copper extraction arein situ leaching or ocean mining.

In situ leaching is a hydrometallurgical pro-cess in which copper is extracted by chemicaldissolution in sulfuric acid. This method is suit-able for low-grade copper ore bodies for whichcustomary mining operations would be uneco-nomical, as well as for the leaching of remenantores from abandoned mines. In some cases, theore body must be broken before leaching byblasting with explosives to increase the surfacearea for chemical reaction.

Ocean mining involves obtaining metallifer-ous raw materials from the deep oceanic zones.Two groups of substances are of interest: deep-sea nodules [43] andmarine ore slimes [44]. Thenodules (manganese nodules; see→Manganeseand Manganese Alloys, Chap. 9.) contain, inaddition to iron oxides, ca. 25 %Mn, 1 %Ni,0.35 %Co and 0.5 % (max. 1.4 %) Cu. Spe-cially equipped ships have collected and liftedthese nodules from depths of 3000 – 5000m;specific metallurgical and chemical methods forprocessing the nodules have been developed inpilot plants. Because of the extremely high ex-penses, large-scale operations of this type havenot yet been undertaken.Marine ore slimes fromthe Red Sea (2200-m depth) average ca. 4 %Zn,1 %Cu, and a little silver. Although methods forprocessing these slimes have been investigated,this resource is not now economically important.

5. Production

Over the years copper production methods havebeen subjected to a continual selection and im-provement process because of the need for (1) in-creased productivity through rationalization, (2)lower energy consumption, (3) increased envi-ronmental protection, (4) increased reliability ofoperation, and (5) improved safety in operation.During this development a number of tendencieshave become apparent:

1) Decrease in the number of process steps2) Preference for continuous processes over

batch processes3) Autogenous operation4) Use of oxygen or oxygen-enriched air5) Tendency toward electrometallurgical meth-

ods6) Increased energy concentration per unit of

volume and time7) Electronic automation, measurement, and

control8) Recovery of sulfur for sale or disposal9) Recovery of valuable byproducts

The selection of a particular productionmethod depends essentially on the type of avail-able raw materials, which is usually ore or con-centrate and on the conditions at the plant loca-tion.

Copper 11

About 80 % of primary copper productioncomes from low-grade or poor sulfide ores. Af-ter enrichment steps, the copper concentratesare usually treated by pyrometallurgical meth-ods. Generally, copper extraction follows the se-quence:

1) Beneficiation by froth flotation of ore to givecopper concentrate

2) Optional partial roasting to obtain oxidizedmaterial or calcines

3) Two-stage pyrometallurgical extractiona) smelting concentrates to matteb) converting matte by oxidation to crude

(converter or blister) copper4) Refining the crude copper, usually in two

stepsa) pyrometallurgically to fire-refined copperb) electrolytically to high-purity electrolytic

copper

Figure 2 [13] illustrates the principal pro-cesses for extracting copper from sulfide ore.

About 15 %, with an increasing trend, of theprimary copper originates from low-grade oxi-dized (oxide) or mixed (oxidized and sulfidic)ores. Such materials are generally treated by hy-drometallurgical methods.

The very few high-grade or rich copper oresstill available can be processed by traditionalsmelting in a shaft furnace. This process is alsoused for recovering copper from secondary ma-terials, such as intermediate products, scrap, andwastes.

Figure 3 [13] illustrates the most importantoperations in copper extraction from various ox-idic ores.

5.1. Beneficiation

Sulfidic copper ores are too dilute for directsmelting. Smelting these materials would re-quire too much energy and very large furnacecapacities. The copper ore coming from themine(0.5 – 1 % Cu) must be concentrated by benefi-ciation. The valuable minerals like chalcopyriteare intergrown with gangue. Therefore, in thefirst step the lumpy ore is crushed and milledinto fine particles (< 100µm) to liberate the in-dividual mineral phases.

Typical equipment for crushing to about20 cm are gyratory and cone crushers. Then wet

grinding in semi-autogenous rod or autogenousball mills takes place. Size classification takes isperformed in cyclones. In the next step of bene-ficiation, valuable minerals and gangue are sep-arated by froth flotation of the ore pulp, whichexploits the different surface properties of thesulfidic copper ore and the gangue [46]. The hy-drophobic sulfide particles become attached tothe air bubbles, which are stirred into the pulp,rise with them to the surface of the pulp, and areskimmed off as a froth of fine concentrate. Thehydrophilic gangue minerals remain in the pulp.Organic reagents with sulfur-containing groupsat their polar end, such as xanthates, are used ascollectors in the flotation process. Additionally,modifiers like hydroxyl ions (pH adjustment)are used to select different sulfide minerals, forexample, chalcopyrite and pyrite. Alcohols areused to stabilize the froth.

To obtain concentrates with highest possiblepurity and recovery rate, the flotation processusually consists of several stages which are con-trolled by expert systems. Various sensors forparticle size, pH, density, and other propertiesare installed. Figure 4 gives an overview of atypical beneficiation process at a concentrator.In the first flotation stage, as much copper aspossible is recovered in a rougher concentrateso that as little as possible goes to the tailings.To increase the copper recovery rate, often thesetailings are leached with sulfuric acid. After re-grinding, the rough concentrate is cleaned in sev-eral flotation steps. After sedimentation in thick-eners and filtration in automatic filter presses orvacuum filters (ceramic disk) the typical cop-per concentrate contains 25 – 35%Cu and about8 % moisture. The moisture content of the con-centrate is a compromise between transportingwater (cost) and avoiding dust generation dur-ing transport. Dewatered concentrates may heatspontaneously or even catch fire; therefore, ap-propriate precautions must be taken [47].

Copper concentrators typically treat up to100 000 t of ore per day.They are locateddirectlyat the mines to achieve low transport costs. Thecopper recovery efficiency is over 90 %. About95%of the ore input goes into the tailings,whichare stored in large dams near the mine and areused for water recycling to the flotation stages.

Separation of special copper ores such asthose containing molybdenite or with high zinc

12 Copper

Figure 2. Principal processes for extracting copper from sulfide ore

Copper 13

Figure 3. Principal processes for extracting copper from oxidic ore [45]

14 Copper

Figure 4. Overview of a typical beneficiation process at a concentrator

or lead content (Canada) is also possible byflota-tion methods. Flotation of non-sulfide copperminerals is rare because these ores are mostlysubjected to hydrometallurgical copper recov-ery, for example, heap leaching. In Zambia andZaire, however, siliceous copper oxide ores arefloated with fatty acid collectors, and dolomiticcopper oxide ores are sulfidizedwith sodium hy-drogensulfide and then floated [48].

5.2. Roasting

Roasting can be used to prepare sulfide concen-trates for subsequent pyrometallurgical or hy-drometallurgical process. Partial roasting underoxidizing conditions may be carried out priorto smelting in reverberatory or electric furnaces.Complete oxidizing or sulfatizing roasting maybe performed before leaching operations, espe-cially if other valuable metals such as cobalt arepresent in the concentrate. Reducing roastingmay be carried out if copper concentrates withvery high contents of impurities such as As areto be smelted.

However, roasting processes are today notvery important for the copper extraction pro-cess. Only a few plants are still operating, forexample Boliden’s Ronnskar Smelter [211] and

Bor Smelter [212]. Since ca. 1975 combinedroasting and matte smelting processes such asflash smelting have been favored because of theirlower energy consumption and process gas han-dling advantages.

The roasting process has several effects:

1) Drying the concentrates2) Oxidizing a part of the iron present3) Controlling the sulfur content4) Partially removing volatile impurities, espe-

cially arsenic5) Preheating the calcined feed with added

fluxes, chiefly silica and limestone

Chemical Reactions. When the moist con-centrates, which contain many impurity ele-ments, are heated, a multitude of chemical reac-tions occur. Because analysis of the many ther-modynamic equilibria is not practical, a few fun-damental systems are usually chosen. The mostimportant is the ternary copper – oxygen – sulfursystem (Fig. 5). The next most important systemis the ternary iron – oxygen – sulfur system be-cause most sulfidic copper ores contain signifi-cant amounts of iron.

Initially, sulfides such as pyrite and chalcopy-rite decompose and generate sulfur vapor, whichreacts with oxygen to form sulfur dioxide:

FeS2 −→ FeS + S(g)

Copper 15

2CuFeS2 −→ Cu2S+ 2 FeS + S(g)

S(g) +O2(g) −→ SO2(g)

The principal reactions, i.e., the formation ofmetal oxides, sulfur trioxide, and metal sulfates,are exothermic.

MS+ 1.5 O2 MO+SO2

SO2 + 0.5 O2 SO3

MO+SO3 MSO4

In addition, there are secondary reactions, suchas the formation of basic sulfates, ferrites (espe-cially magnetite), and silicates, the last provid-ing most of the slag in the subsequent smelting:

MO+MSO4 −→ MO ·MSO4

MO+Fe2O3 −→ MFe2O4

FeO+Fe2O3 −→ Fe3O4

MO+SiO2 −→ MSiO3

Representative reductive roasting reactions are:

FeS2 −→ FeS + S(g)

8 FeAsS −→ 4 FeAs + 4 FeS +As4S4(g)

Figure 5. Partial phase diagram of the ternary Cu –O–Ssystem [51]

Methods. There are several important roast-ing methods; all involve oxidation at an elevatedtemperature, generally between 500 and 750 C:

1) Partial (oxidizing) roasting is the conven-tional way of extracting copper from sul-fide concentrates. At 700 – 750 C, the de-gree of roasting is determined by controllingthe access of air. A predetermined amount ofsulfur (30 – 50 % is removed, and only partof the iron sulfide is oxidized. The coppersulfide is relatively unchanged. These con-ditions are important for the formation of asuitable matte.

2) Total, or dead, roasting is occasionally usedfor complete oxidation of all sulfides for asubsequent reduction process or for specialhydrometallurgical operations.

3) Sulfatizing roasting is carried out at550 – 650 C to form sulfates. This methodyields calcines well-suited for hydrometal-lurgical treatment.

Roasters. Industrial roasting is done in twotypes of roasters: fluidized-bed and multiple-hearth roasters. Both are continuously operatedprocesses.

Oxidizing roasting is usually carried out influidized-bed roasters with short residence timesin the range of seconds and high production ratesup to 50 t of moist concentrate per hour. Theoxidation reactions supply most of the requiredheat. About 30 – 50 % of the incoming sulfideis oxidized to SO2 by using slightly oxygen en-riched air (up to 30 % O2). The off-gas is richin SO2(6 – 12 %) and suitable for conversion tosulfuric acid. The hot calcine is usually sent toreverberatory or electric furnace. The advantageof a roaster in front of a smelting furnace isthe lower energy requirement of the smeltingfurnace and the higher matte grade. Examplesof fluidized-bed roasters are Boliden Ronnskar(Sweden), Bor Smelter (Yugoslavia) in front ofpyrometallurgical copper extraction, and Cham-bishi (Zambia) ahead of a leaching plant.

Reductive roasting is usually carried out inmultiple-hearth furnaces because of the long res-idence time (several hours) and the precise con-trol of temperature and gas composition on eachhearth. These roasters have lower productionrates and are fired by natural gas burners. An ex-ample for the reducing process is the treatment

16 Copper

of El Indio (Chile) concentrate, which contains8%As and about 22%Cu.About 97%of the ar-senic is removed during the roasting process atabout 650 – 720 C [213]. In former times themultiple-hearth roasters (Herreshoff furnaces)were widely used for the extraction of copperfrom concentrates.

5.3. Pyrometallurgical Principles

Smelting of unroasted or partially roasted sul-fide ore concentrates produces two immisciblemolten phases: a heavier sulfide phase contain-ing most of the copper, the matte, and an oxidephase, the slag. In most copper extraction pro-cesses, matte is an intermediate.

5.3.1. Behavior of the Components

The most important equilibrium in copper mattesmelting is that between the oxides and sulfidesof copper and iron:

Cu2O+FeS Cu2S+FeO

Iron(II) oxide [1345-25-1] reactswith added sil-ica flux to form fayalite [13918-37-1], a ferroussilicate that is the main component of slag:

2 FeO+SiO2 −→ Fe2SiO4

Liquid iron sulfide [1317-37-9] reduces higheriron oxides to iron(II) oxide:

3 Fe2O3 +FeS −→ 7 FeO+SO2(g)

3 Fe3O4 +FeS −→ 10 FeO+SO2(g)

The second reaction serves to removemagnetite[1309-38-2], which complicates furnace opera-tions because of its highmelting point (1590 C)[54].

The pyrometallurgical production of copperfrom sulfide ore concentrates may be consideredas a rough separation of the three main elementsas crude copper, iron(II) silicate slag, and sulfurdioxide. About 20 accompanying elements mustbe removed from the copper by subsequent refin-ing. Table 8 shows the distribution of importantimpurities amongmatte, slag, and flue dust. Pre-cious metals, such as silver, gold, platinum, and

palladium, collect almost entirely in the matte,whereas calcium,magnesium, and aluminum gointo the slag.

Table 8. Average percentage distribution of the accompanying ele-ments in copper smelting, p. 591[20]

Element Matte Slag Fluedust

Arsenic 35 55 10Antimony 30 55 15Bismuth 10 10 80Selenium 40 – 60Tellurium 40 – 60Nickel 98 2 –Cobalt 95 5 –Lead 30 10 60Zinc 40 50 10Tin 10 50 40Silver and gold 99 1 –

5.3.2. Matte

The ternary Cu – Fe – S system is discussed indetail in the literature [55–57]. Figure 6 showsthe composition of the pyrometallurgically im-portant copper mattes and the liquid-phase im-miscibility gap between matte and the metallicphase. In the liquid state, copper matte is essen-tially a homogeneous mixture of copper(I) andiron(II) sulfides: the pseudobinary Cu2S – FeSsystem.

Arsenides and antimonides are soluble inmolten matte, but their solubility decreases withan increasing percentage of copper in the matte.Accordingly, when the arsenic concentration ishigh, a special phase, the so-called speiss, canseparate. It is produced under reducing condi-tions in the blast or electric furnace, and itsdecomposition is complicated (→Arsenic andArsenic Compounds, →Antimony and Anti-mony Compounds).

Compositions of several copper mat-tes are shown in the partial diagramCu2S – FeS – (Fe3O4 +FeO) (Fig. 7), which is asection of the quaternaryCu – Fe –O – S system.The density of solid copper mattes ranges bet-ween 4.8 g/cm3 (FeS) and 5.8 g/cm3 (Cu2S);liquid mattes have the following densities:4.1 g/cm3 (30wt %Cu, 40wt %Fe, 30wt %S),4.6 g/cm3 (50wt % Cu, 24wt %Fe, 26wt %S),and 5.2 g/cm3 (80wt %Cu, 20wt %S).

Copper 17

Figure 6.Ternary Cu – Fe – S diagram showing coppermat-tes and the miscibility gap [55]

Figure 7. Partial ternary Cu2S – FeS1.08 – (Fe3O4 +FeO)diagram [58] showing mattes from various processesReverberatory furnace;Flash smelting furnace;Elec-tric furnace; • Blast furnace; Converter

Table 9. Composition (wt %) of typical copper smelter slags [64]

Component Reverberatory Flash Noranda Peirce – Smithfurnace furnace reactor converter

Copper 0.4 – 0.6 1 – 2 8 – 10 2 – 8Iron (total) 35 40 35 50Silica 38 30 21 25Magnetite 7 – 12 13 25 – 29 20 – 25Ratio of Fe to SiO2 0.92 1.33 1.67 2.0

5.3.3. Slags

Slags from copper matte smelting contain30 – 40 % iron in the form of oxides and aboutthe same percentage of silica (SiO2 ), mostlyas iron(II) silicate. Such slags can be consid-ered as complex oxides in the CaO – FeO –SiO2system [59] or, because of the relatively lowCaO content of most slags, in the partial dia-gram FeO–Fe2O3– SiO2 [60] (Fig. 8). Ternarysystems of these and other pertinent oxide sys-tems are found in the literature [61], [62]. Ta-ble 9 shows the general composition of someslag types. Important properties of copper slagsystems are compiled in [63].

Figure 8. Ternary FeO – Fe2O3– SiO2 diagram [60]

The objectives of matte smelting are toachieve a rapid, complete separation of matte

18 Copper

and slag, the two immiscible phases, and a min-imal copper content in the slag. The differingproperties of slag and matte affect this separa-tion:

1) the low, narrow melting interval of slag2) the low density of liquid slag (ca.

3.1 – 3.6 g/cm3) and the difference in den-sity between molten matte and slag of ca.1 g/cm3

3) the low viscosity and high surface tension ofthe slag

The ratio of the weight percent of copper inmatte to that of copper in slag should be bet-ween 50 and 100. High matte grades generallycause high copper losses in slag. Such lossesdepend on the mass ratio of slag to copper pro-duced,which is usually between 2 and 3. Copperin slags occurs in various forms, including sus-pended matte, dissolved copper(I) sulfide, andslagged copper(I) oxide, partially as a silicate,which is typical of nonequilibrium processing.

Slags containing < 0.8 % copper are soldas products with properties similar to those ofnatural basalt (crystalline) or obsidian (amor-phous) or discarded as waste. When liquid slagis cooled slowly, it forms a dense, hard, crys-talline product that can be used as a large-sizefill for riverbank protection or dike constructionand as amedium-size crushedfill for roadbeds orrailway ballast. Quick solidification, by pouringmolten slag into water, gives amorphous granu-lated slag, an excellent abrasive that has partiallysupplanted quartz sand. Ground granulated slagis used as a trace element fertilizer because ofits copper and other nonferrous metals.

Most of the newer copper smelting pro-cesses produce high-grade mattes, and the shortresidence time of the materials in the reac-tion chamber results in an incomplete approachto chemical equilibrium. Both factors lead tohigh amounts of copper in the slag, generally>1wt %. Such slags must be treated by specialmethods for copper recovery (Section 5.5.1).

5.3.4. Oxidizing Smelting Processes

Nearly all pyrometallurgical copper processesare based on the principle of partial oxidationof the sulfide ore concentrates. Methods based

on the total oxidation of sulfide ores with subse-quent reduction tometal, avoiding the formationof copper matte, are used only rarely because ofhigh fuel consumption, formation of copper-richslags, andproduction of crude copperwith a highlevel of impurities.

Prior to the 1960s, the most important way ofproducing copper was roasting sulfide concen-trates, smelting the calcines in reverberatory fur-naces, and converting thematte inPeirce – Smithconverters. Since that time, the modern flashsmelting processwith subsequent converting hasbecome predominant. Figure 9 shows the flowsheet of a modern copper smelter, from concen-trate to pure cathode copper, including the useof oxygen, recovery of waste heat, and environ-mental protection. Table 10 compares the impor-tant stages and processes of copper production,showing the range of the matte composition foreach process.

Figure 9. Typical flow sheet for pyrometallurgical copperproduction from ore concentrates [65]

5.3.5. Proposals

Numerous laboratory experiments and pilot-plant runs have been carried out to developsmelting methods based on elements other than

Copper 19Table10.S

urveyof

pyrometallurgicalprocessesforcopper

productio

n[66]

20 Copper

oxygen. Two lines of development have dom-inated, reduction with hydrogen and chlorina-tion, but without leading to commercialization.

Reduction. A potential process involves thereduction of chalcopyrite [70]:

2CuFeS2 + 3H2 + 3CaO −→ Cu2S+ 2 Fe + 3CaS + 3H2O

CuFeS2 + 2H2 + 2CaO −→ Cu+Fe + 2CaS + 2H2O

The reduction by hydrogen is endothermic,but the overall reaction with calcium oxide isexothermic. A similar proposal [71] is based onthe reaction of a metal sulfide with steam in thepresence of calcium oxide:

Chlorination. The reactions of chalcopyritewith chlorine are also of interest [72]:

>500 C:

CuFeS2 + 2Cl2 −→ CuCl2 +FeCl2 + 2 S

<500 C:

CuFeS2 + 3.5Cl2 −→ CuCl2 +FeCl3 +S2Cl2

followed by electrolysis of the molten copper(I) chloride.

The recently proposed thermoelectron chlo-rination process is another variation [73]:

CuFeS2 + 2Cl2 −→ CuCl + FeCl3 + 2 S

Electrolysis. Another approach, to avoidconverting, proposed the electrolysis of moltencopper matte [74].

5.4. Traditional Bath Smelting

At the end of the Middle Ages, copper was pro-duced by the German or Swedish smelting pro-cess that involved roasting reduction with upto seven process steps in small shaft furnaces.

Around 1700, reverberatory furnaces were con-structed inwhich the orewas processed by roast-ing, the so-calledEnglish orWelsh copper smelt-ing process, originally with ten process steps.

The large blast and reverberatory furnaces ofthe 1900s were derived from these principles.Later, the electric furnace for matte smeltingwas developed. Newer processes are the Isas-melt/Ausmelt/Csiromelt (furnace with verticalblowing lance), the Noranda and CMT/Tenientereactors (developed from converters), the Rus-sian Vanyukov, and the Chinese Bayia process.

5.4.1. Blast Furnace Smelting

The blast or shaft furnace is well-suited forsmelting high-grade, lumpy copper ore. If onlyfine concentrates are available, they must firstbe agglomerated by briquetting, pelletizing, orsintering. Because of this additional step and itsoverall low efficiency, the blast furnace lost itsimportance for primary copper production and iscurrently used in only a few places, for example,Glogow in Poland.

Smaller types of blast furnace, however, areused to process such copper-containing materi-als as intermediate products (e.g., cement cop-per or copper(I) oxide precipitates), reverts (e.g.,converter slag, refining slag, or flue dusts), andcopper-alloy scrap.

The construction of the furnace is basicallyrelated to that of the iron blast furnace, butthere are considerable differences in design, es-pecially in size and shape: the copper blast fur-nace is lower and smaller, and its cross sectionis rectangular. Developments adopted from thesteel industry include use of preheated air (hotblast), oxygen-enriched air, and injection of liq-uid fuels.

The furnace is charged with alternate addi-tions of mixture (copper-containing materialsand accessory fluxes such as silica, limestone,and dolomite) and coke (which serves as bothfuel and reducing agent). There are three zonesin the furnace:

1) In the heating zone (the uppermost), waterevaporates and less stable substances disso-ciate.

2) In the reduction zone, heterogeneous reac-tions between gases and the solid charge takeplace.

Copper 21

3) In the smelting zone, liquid phases react.

The usual mode of operation is reduc-ing smelting, which yields two main prod-ucts. Sulfide ores are used to produce amatte (40 – 50wt %Cu) and a disposal slag(ca. 0.5wt %Cu). In contrast, oxide ores areprocessed directly to impure black copper(95wt %Cu) and to a copper-rich slag. Thetwo ore types can be smelted together to pro-duce matte and a slag with low copper content.Another product is top gas, which contains fluedust. Ores that contain high concentrations ofarsenic and antimony also form speiss, which isdifficult to decompose.

In Poland (KGHM Polska Meidz S.A.Smelters in Glogow I and Legnica) the blast-furnace technology is well adapted to Pol-ish copper concentrates, which contain about20 – 30 % Cu like normal chalcopyrites but also5 – 10 % of organic carbon and only 9 – 12 % S.Also these concentrates have relatively high lead(up to 2.5 %) and arsenic (up to 0.3 %) content.The organic carbon compounds provide about40 – 60%of the process energy; the rest is addedby coke. The matte has about 58 – 63 % Cu and3 – 6 % Pb. The slag contains less than 0.5 %Cu. The off-gas from the blast furnaces (threein each plant) is mixed with the converter gases(Hoboken Converter). It contains 7 – 10 % SO2and is sent to sulfuric acid production. The pro-duction figures are 80 000 t/a converter copperin Legnica, and 200 000 t/a in Glogow I Smelter[214].

5.4.2. Reverberatory Furnace Smelting

The reverberatory furnace dominated coppermatte smelting for much of the 1900s, because itwas an excellent process for smelting fine con-centrate from flotation. It is a fossil fuel firedhearth furnace for smelting concentrate and pro-ducing copper matte. The reverbs began to de-cline in the 1970s with the adoption of environ-mentally and energetically superior processeslike flash smelting. Probably the last one waserected in 1976 in Sar Chemesh, Iran. In 1980about 100 reverbs were in operation, but in 1994the number had decreased to about 25. As shownin Figure 10 it is a rectangular furnace up to10mwide and 35m long with internal brick lin-ing. Throughputs of up to 1100 t/d concentrate

or a mixture of concentrate and calcine couldbe processed. The charge is passed into the fur-nace near the burners through the roof or lateralopenings. As fossil fuel, pulverized coal, heavyoil, or natural gas is used. Normally the burneris located in the front wall of the furnace. Theatmosphere is slightly oxidizing, and the maxi-mumflame temperature is up to 1500 C.Duringthe 1980s oxygen – fuel burners have been set inthe roof to fire downwards directly on the top ofthe bath. This increases the smelting rate by upto 40 % and the energy efficiency to about 50 %[76], [215]. Another invention was the sprinklerburner for feeding concentrate, coal, and fluxfrom the top of the reverb [77].

The back half of the furnace is the settlingzone. A matte grade of between 35 and 60 % Cuis produced, depending on whether concentrateor calcine is fed. The slag has low copper con-tent (< 1%). Normally, the converter slag is alsofed back to the reverb. Sometimes problemswithsolid magnetite accretions in the furnace occur.

The off-gas has a temperature of about1250 C. It is diluted by the combustion air andcontains only about 1 – 2 % SO2. This is toolow for efficient SO2 capture such as sulfuricacid production by the contact process. Improve-ments have been achieved by using higher oxy-gen enrichment of the burners, but the off-gasis still too dilute, and therefore this furnace isunsuitable for many regions in the world be-cause of environmental problems. Another dis-advantage is the very high energy consumption.The reverb has the highest energy consumptionof all copper matte smelting processes. A goodpresent example for operating reverbs is the IloSmelter in Peru (Table 11). There two reverbssmelt about 2000 t/d of concentrate, producing amatte containing 35%copper for the subsequentPeirce – Smith converters. In the latest smelterenlargement one reverb was replaced by an Te-niente converter (CMT), and a sulfuric acid plantfor partial SO2 recovery was built.

5.4.3. Electric Furnace Smelting

In regions where relatively cheap electricalpower was available, electric furnaces werebuilt. The Scandinavian countries were the firstto perform electricmatte smelting: 1929 at Sulit-jelma, Norway; 1938 at Imatra, Finland, and

22 Copper

Figure 10. Schematic longitudinal and cross-section views through a reverberatory furnace [45]

1949 at Ronnskar, Sweden, which is still op-erating.

Electric furnaces have the rectangular groundplan and the dimensions of larger reverbera-tory furnaces. Along the centerline, up to sixSoderberg continuous self-baking electrodes areused with alternating current (Table 11).

Table 11.Examples of traditional reverberatory and electric furnaces

Boliden RonnskarSmelter

Southern Peru IloSmelter

Type electric furnace reverbatoryInside dimensionsl×w× h, m

24× 6× 3 36× 10× 3.4

Electrodes/burners 6 8Smelting capacity, t/d 950 (hot calcine) 1000 (wet

concentrate)Matte grade, % Cu 51 35Off-gas volume, m3/h(STP)

25 000 100 000

SO2 content, % 4.5 1.2Energy consumption 300 kW/t concentrate 194 kg oil/t

concentrate

The smelting process in such units is similarto the operation in reverbs, but the concentrate isusually dried and roasted before charging to in-crease the smelting capacity. The converter slagis returned to the furnace. The composition ofmatte and slag resembles that of the reverb prod-ucts, but the content of magnetite is lower. Nofuel is burned, and the volume of waste gasesand the quantity of flue dust are small. The SO2content of thewaste gas can be 10%.The off-gasalso contains carbon oxides from the consump-tion of the graphite electrodes.

A significant difference between the electricfurnace and the reverb is the inversion of the tem-perature gradient in the furnace cavity. In the re-verb, the combustion gases have the highest tem-perature, whereas in electric matte smelting, thewaste gases are ca. 500 C cooler than the slag

phase, which is heated by the electric energy.Accordingly, in the electric furnace, cheap re-fractories are sufficient for lining thewalls abovethe slag zone and the roof; only a common archis required.

The temperatures of both molten phases de-pend on the submergence of the electrodes, andthe required heat is controlled by the electricalpower supplied. Heat transfer takes place chieflyby convection, which causes intense circulationin the molten bath.

The current is divided into two partial cir-cuits, through slag and through matte. The dif-ference in conductivity is great, slag : matte ra-tio of 1 : 102 to 1 : 103 ; therefore, the depth ofimmersion of the electrodes into the liquid slagmust be precisely controlled. As the electrodesimmerse deeper into the slag,more current flowsthrough the matte. If they touch the matte layer,a short circuit occurs. These considerations leadto an approximate relation between the electri-cal power input and the depth of the slag layer:6000 kVA and 0.5m, 30 000 kW and 1.0m, and50 000 kW and 1.5m.

The smelting capacity of electrical furnacesis higher than that of reverbs.

BrixleggProcess. Lurgi developed and prac-ticed at Brixlegg, Tirol, Austria, a modificationof the old roasting reactionprocess.Nearly dead-roasted concentrates are reduced in batches bycoal in a small special circular electric furnace(2500 kVA, 5-mdiameter). The crude copper av-eraged only 95wt %Cu, and the operation hasbeen discontinued.

Copper 23

5.4.4. Isasmelt Furnace

The Isasmelt furnace and the closely relatedAusmelt furnace were both developed in the1980s on the basis ofwork conducted byCSIRO,Australia [216]. The process consists of a tallcylindrical furnace (small diameter but largeheight, internally brick lined) equipped with avertical blowing lance. The lance is submergedinto the bath (slag) and blows oxygen-enrichedair and/or fossil fuel (natural gas). The moistconcentrate and flux material is pelletized andfed through the roof of the furnace onto the bath.Most of the energy requirement comes from thereaction of the concentrate. The lance is madefrom stainless steel. It is cooled by swirling theair in the annulus between the pipes so that aprotective layer freezes on that part of the lancewhich is immersed in the slag bath (0.3m) [217].The lance has to be replaced from time to time,for example, once a week. In the continuous Isasprocess, an emulsion of matte and slag is pro-duced, which is tapped periodically to a electricor fuel-fired settling furnace for matte/slag sep-aration. The off-gas is high in SO2. For coolingit passes through a waste-heat boiler.

Two Isasmelter for copper matte are running(Mount Isa, Australia; Cypus Miami, Arizona).Figure 11 shows the Cyprus furnace schemati-cally. About 90 t/h of copper concentrate is pro-cessed, and about 45 t/h of matte is produced(58%Cu). The off-gas contains about 35%SO2directly behind the furnace and is then diluted byair to 8 – 9 %. The Isasmelt process has replacedthe electric (reverberatory) furnace, which isnow used for matte/slag settling. Smaller Isas-melt or Ausmelt furnaces have also been pro-posed or built for secondary copper smelters andprimary smelters like in China. It is possible torun the process batchwise, smelting and con-verting the material in the furnace for producingblister copper. Also Isasmelt furnaces have beenbuilt for lead refining.

5.4.5. Noranda Process

The Noranda process was initially constructedas a continuous smelting and converting processwhich produced blister copper from copper con-centrate. The first reactor was built in 1973 atHorne smelter [105]. This direct-to-blister pro-cess operated from 1973 to 1975 [218]. It was

switched to high-grade copper matte smelting(as it now still operates) because of excessiveimpurity levels in the copper anodes and to in-crease the smelting rate.

Figure 11. Isasmelt furnace

The Noranda furnace [219] is a horizontalsteel barrel with an inner brick lining (Figure 12.The diameter is about 5m and it is about 20mlong. The process runs continuously. At one endof the reactor, pelletized wet concentrate, coal,flux, revert materials, and scrap are thrown intothe furnace and on to the top of the molten bathby a high-speed slinger belt. Feeding fine con-centrate through the tuyeres is also possible. Thefeedmaterial is absorbed andmelted in the liquidmatte/slag bath.On the side there are 20 – 40 tuy-eres in the cylindrical part of the vessel. Oxygen-enriched air (ca. 40%O2) is blown through thesetuyeres into the matte phase. The tuyeres have tobe punched periodically, as in a Peirce – Smithconverter. A layer of matte and slag must alwaysbe present in the furnace.

The oxidation reaction of the sulfides (mostlyiron sulfide) and the added coal provide the pro-cess energy. Matte is tapped at the bottom ofthe vessel. Slag is periodically tapped at the

24 Copper

other end of the cylindrical furnace where thereare no tuyeres and the bath is quiet enough formatte and slag settling. The matte usually con-tains 70 – 75 % Cu (white metal) and is sent toa subsequent converter. This operation gives alarge heat evolution from Fe and S oxidation inthe Noranda furnace. This avoids excessive re-tention of impurities like Sb in the matte. Theslag contains about 3 – 7 % copper and goes tofurther treatment like electric furnace or solidifi-cation/comminution/flotation. The off-gas con-tains about 15 – 20 % SO2. It passes through thelarge mouth at the top of the reactor into a hoodand is sent to cooling, cleaning, and sulfuric acidproduction. In case of emergencies the furnacecan be rolled out so that the tuyeres have a secureposition above the molten bath.

Figure 12. Noranda furnace

The Noranda process is well suited for smelt-ing scrap and residues, ashes, precious metalsscrap, telephone scrap, and computer/electronicscrap. Due to the high temperature and smeltingintensity, potentially harmful organics are oxi-dized to CO2 and H2O. The productivity can beincreased by using higher oxygen enrichment.Then less fossil fuel is needed and the amountof off-gas decreases.

5.4.6. CMT/Teniente Process

The Caletones Matte Treatment (CMT) or socalled Teniente process [220], [221] is similar tothe Noranda process. The important differencesare that no coal is included in the charge andmolten low-grade copper matte is periodicallyadded as fuel. Therefore, the Teniente furnace ismostly regarded as a converter.

The process was developed at TenienteSmelter by adding dry concentrate to a

Peirce – Smith converter for increasing exist-ing smelter capacity. The Teniente process hasbeen widely adopted in Latin America. Likethe Noranda reactor, it is a robust technologybut has still not demonstrated that it can meetthe highest environmental standards. The fur-naces are horizontal cylinders, 4 – 5m in diam-eter and 14 – 22m long (Figure 13). Inside, theconverter is brick-lined, and it has 30 – 50 tuy-eres for blowing oxygen-enriched air (30 – 32%O2). Solid feed such as wet concentrate, flux,or recycled materials are blown continuouslythrough a high-pressure gun (Garr gun) on thetop of the molten bath. Dry concentrate is in-jected additionally through four or five tuyeresin the bath.Moltenmatte (35 – 40%Cu) comingfrom a reverb is added periodically through themouth of the converter. The temperature in theconverter is controlled by the ratio of wet (low-ers temperature) and dry (increases temperature)feed material.

High-gradematte (72 – 77%Cu) is producedand tapped periodically to the subsequent con-verting step. Slag with 4 – 7 % Cu is tapped forfurther processing. The off-gas is collected by awater-cooled hood above the mouth of the con-verter. The off-gas is mixed with air because thehood is not tight to the converters mouth. There-fore, the SO2 content is only about 10 %. Thegas is cooled with water sprays, the flue dust isstripped and then the gas is sent to sulfuric acidproduction (or in some cases to the stack).

With its molten-matte requirement, the CMTis dependent on the operation of a reverbera-tory furnace with all the associated environmen-tal problems (SO2 emission). To overcome thisreliance, some possibilities have been tested,all of which increase the energy in the vessel:increased oxygen enrichment of the blast, in-creased injection of dry concentrate through thetuyeres, and burning small amounts of fossilfuel. Today there are CMTs which do not needlow-grade molten matte.

5.4.7. Vanyukov Process

The Vanyukov process was developed in the1970s in Norilsk, Russia. It is similar to the No-randa process, but the furnace is rectangularwithtuyeres on both sides, blowing highly enrichedair ( 70 – 90 % O2) in the slag zone (Figure 14).

Copper 25

Figure 13. Teniente converter (CMT)

Moist copper concentrate is charged throughports in the roof of the furnace. The process pro-duces matte (ca. 50 % Cu), slag (1.5 – 2 % Cu)and off-gas (ca. 30 % SO2). The furnace oper-ates continuously, and the matte and slag leavethe furnace continuously through two siphons.The matte is sent to converting process, and theslag is cleaned in an electric furnace [222], [223].

In contrast to the Noranda furnace theVanyukov reactor is stationary. The tuyeres cannot be tilted out of the bath in emergency situ-ations. The advantage of the stationary systemis the simple and very efficient SO2 capture. SixVanyukov furnaces are running in the CIS. Witha relatively small furnace (10m long, 2m wide,and 6mhigh) about 1500 to 2000 t/d concentratecan be processed.

All the submerged-tuyere processes like No-randa, CMT, and Vanyukov have relatively shortcampaign lifetimes (one year) because of higherosion in the tuyere zone.

5.4.8. Baiyin Process

At theBaiyinCopperSmelter inChina, a processsimilar to the Noranda furnace was developedin the 1980s [78]. Concentrate is added on thetop of the molten bath, and air/oxygen is blownthrough submerged tuyeres into the melt. TheBaiyin furnace has a wall inside to divide the re-action and the settling zone. In China there areat least two furnaces in operation, but few dataare available.

5.5. Autogenous Smelting

Autogenous smelting involves the use of com-bustion heat generated by reactions of the feedin an oxidizing atmosphere in which the sulfideconcentrate acts partly as a fuel. The formerlyseparate steps of roasting and smelting are com-bined into a roast – smelting process. The spatialand temporal coupling of exothermic and en-dothermic reactions leads to an economical pro-cess, but the sensible heat of nitrogen in the aircauses a deficit in the heat balance. In practice,various measures must be taken:

1) Increasing the oxygen content of the com-bustion air and even using pure oxygen

2) Preheating combustion air with waste heat orin a preheater

3) Combustion of natural gas, fuel oil, or pul-verized coal in supplementary burners

To achieve autogenous operation and preventagglomeration of the feed, the moisture in theconcentrate must be removed by drying beforecharging. The quantity of added fluxes is mini-mized as far as practical to save energy.

Because the residence time of the sulfide par-ticles in the reaction chamber is only a fewseconds, kinetic conditions predominate overthe thermodynamic equilibrium. The reactantsform a heterogeneous system, with the feed sus-pended in the gas flow, thus the term smelting insuspension.

These processes have several advantages:

1) High rate of reaction, increasing the produc-tion rate

2) Energy savings

26 Copper

Figure 14. Vanyukov furnace

3) Low volume of off-gas and correspondinglyhigh concentration of SO2 and low quantityof flue dust, if oxygen is used

However, a typical disadvantage is the high cop-per content of the slags, and the relatively highflue dust content in the off-gas, which can causeproblem in the waste-heat boiler.

5.5.1. Outokumpu Flash Smelting

After a preliminary test in 1946, the first full-scale flash smelting furnace started operation in1949 at Outokumpu Oy, Harjavalta (Finland).Flash smelting has been themostwidely adoptedcoppermatte smelting process since 1970. Morethan 40 furnaces have been installed to replacereverberatory furnaces or at new smelting opera-tions. Nowadays the Outokumpu-type smeltersaccount for more than 50 % of world primarysmelter capacity.

Furnace Design. The flash furnace can beenvisioned as a combination of a reverb witha roaster in which smelting takes place simul-taneously. Figure 15 shows a Outokumpu flashfurnace designed in 1988 for smeltingmore than3000 t/d of copper concentrate – a productionrate several times higher than those of earlierfurnaces. The current trend is towards even big-ger units. The furnace is 24m long, 7.5m wide,and 2m high in the settling area. The hearth orsettler is equipped with oil or natural gas burn-ers.

The reaction shaft is 6m in diameter and 7min height. It is equipped with one concentrateburner. The gas off-take shaft is 9m high andhas a diameter of 5m. (All dimensions insidethe furnace, which is lined with chrome magne-sia bricks.) The furnace has eight matte tapholes

and four slag tapholes. The reaction shaft andlarge areas of the settler and the gas off take arewater-cooled to prevent overheating and weak-ening of the furnace structure. The reaction shaftof many furnaces is cooled by cascading wateroutside the steel shell. The other areas are usu-ally cooled by special copper cooling jackets,which are built into the furnace walls [224].

Figure 15. Outokumpu flash smelting furnace

An important feature of the flash smelter isthe concentrate burner, which suspends the feedparticles in the oxygen-enriched air in the hotreaction shaft. The aim is to ensure a rapid, uni-form (over the shaft, to prevent hot spots), and ef-ficient concentrate oxidation and generate min-imum flue dust. Older flash smelters have fourburners, but nowadays a single burner is usedbecause it is easier to control the burner param-eters. The burner (Fig. 16) consists of two con-centric pipes: a central pipe for the solid feedand an annulus for the oxygen-enriched blast. Inthe inner pipe a distributor cone is installed tooptimize the particle distribution [225]. The gasvelocity out of the burner is up to 200m/s.

Copper 27

Figure 16. Schematic of a flash furnace concentrate burner

Flash furnaces need auxiliary equipment likesolids dryers (rotary, flash, or steam dryers), de-livery systems, oxygen plant, blower and pre-heater systems, waste-heat boiler, dust recoveryand recycling system, slag-treatment furnace,and sulfuric acid plant for treating the highlySO2 enriched off-gas.

Operation. Various sorts of concentrate anda few flux materials (mainly silica) are blendedand then dried before feeding into the flashsmelter. Recyclematerials such as flue dust fromthe waste-heat boiler and the converters or oth-ers are also fed to the flash furnace. The processoperates continuously 24 h a day, and campaignsare run for as long as possible with out interrup-tion.

The feed mixture is suspended at the top ofthe reaction shaft, and sulfur and iron are oxi-dized partly. The reactions are strongly exother-mic, and the particles melt to form copper matteand slag. At oxygen enrichments of between 45and 80 % (depending on concentrate composi-tion and preheated blast), the process runs au-tothermally. Less oxygen enrichment requiresfossil fuel combustion in the reaction shaft.The temperature in the reaction shaft is about1250 – 1300 C. The optimal matte concentra-tion is between 60 and 65%Cu. Higher Cu con-

centrations the in the matte are also possible, forexample, before a flash converting process.

The matte is tapped as needed by the sub-sequent converting step. Slag is tapped nearlycontinuously. The slag contains 1 – 2 % of cop-per, mainly in matte droplets. The slag tem-perature is about 1230 – 1250 C; the matte isabout 30 – 50 C cooler. Typical operating dataare shown in Table 12.

Table 12. Operating data for Outokumpu flash furnaces

NorddeutscheAffinerie, Hamburg,Germany

BHP Magma Metals,San Manuel, Arizona

Furnacecommissioning date

1972 1988

Size, inside brick, mhearth w× l× h

6× 20× 3 7.5× 24× 2

reaction shaft d× h 6× 9 6× 6.9Concentratethroughput, t/d

2000 (30 – 34 % Cu) 3000 (30 % Cu)

Blast temperature,C

400 – 420 25

Oxygen enrichment,%

45 – 55 60 – 80

Blast flow-rate, m3/h(STP)

25 000 51 000

Matte grade, % Cu 63 63Off-gas volume,m3/h (STP)

55 000 55 000

SO2Content 25 – 30 35Recycling ofconverter slag

in flash furnace flotation

Slag treatment electric furnace flotationFossil fuel input inshaft

0 0

Fossil fuel for settlerburner

5m3/h (STP)natural gas/t conc.

15 – 20 L oil/t conc.

Slag Cleaning. In practice two methods areused: froth flotation and treatment in smallelectric slag furnaces. The decision for one orthe other is determined by economics. In frothflotation the slag concentrate is returned to thefeed, and the tailings are discarded. The secondmethod involves the reduction of flash furnaceslags in an electric furnace, which yields a matteand slag that is suitable for sale.

Off-gas and Flue Dust. The waste gases areseparated frompart of thefluedust at ca. 1300 Cin the off-take shaft and pass through a waste-heat boiler for generating steam, and subse-quently to an electrostatic (Cottrell) precipitatorfor separating the mass of flue dust, which is re-cycled to the feed. The precleaned off-gas withSO2 content of >20 % is usually processed tosulfuric acid.

28 Copper

The flue dust consists chiefly of sulfates andbasic sulfates of copper, lead, and zinc, as wellas some volatile compounds of arsenic, anti-mony, bismuth, and selenium. Repeated recy-cling makes it possible to enrich selected ele-ments for later extraction. The quantity of fluedust is between 4 and 10 % of the input.

Process Control. Fivemain parameters haveto be controlled and adjusted carefully to obtainthe highest smelting rate of concentrate, constantmatte composition, specified slag compositionand furnace temperatures, and minimum energyconsumption:

1) Concentrate feed rate2) Flux and recycled materials feed rate3) Blast input and temperature4) Oxygen enrichment of the blast5) Fossil fuel combustion

In the flash smelters constructed between1980 and 1999 most of these parameters aremeasured and adjusted automatically.

Special Operations. There are several vari-ations on theOutokumpu flash smelting process.

Two Outokumpu copper smelters (Tamano,Japan and Pasar, Philippines) have inserted car-bon electrodes in the settler. The electrodes aresubmerged in the slag layer to heat the slag withelectric power. The system was installed to re-cover Cu from the slag and reduce magnetitewithout using an additional slag-cleaning fur-nace [80], [226]. The system at Tamano Smelterceased operation in 1988 because the opera-tors found that combustion of coke – oilmixturesalso gives low copper content in the slag (0.6 %)[227].

Occasionally molten converter slag is recy-cled through the flash furnace like in a reverb. Itis added from ladles through a spoon and laundersystem high on the flash furnace [228].

At two smelters in the world Outokumpuflash furnaces are used to smelt blister directlyfrom concentrate (see Section 5.8).

Trends. The diversity of modified furnaceconstructions and operating methods shows theadaptability of the Outokumpu process to dif-ferent raw materials and smelter locations. Asincreasing oxygen content has been used in the

blast air, the process has approached the princi-ple of INCO flash smelting. On the other hand,newer developments aim at the production ofblister copper or white metal (75 – 80wt %Cu)from concentrates in only one vessel, therebyapproaching continuous smelting and convert-ing as in the Noranda process.

5.5.2. Inco Flash Smelting

Inco Metal Co. was the first company in thenonferrous metal industry to use commerciallypure oxygen for autogenous flash smelting. Af-ter tests in the late 1940s, two smaller furnaceswent into operation in 1953 [82]. Today a totalof five Inco furnaces are operating worldwide(including Ni/Cu and one older plant in Uzbek-istan [83]). Since 1990 no Inco furnace has beenbuilt.

Figure 17. Inco flash smelting furnace

Furnace Design. In contrast to Outokumpu-type furnaces the Inco process uses pure com-mercial oxygen (95 – 98 %) instead of enrichedair. The oxygen blast and concentrate are fedthrough the burners (generally two) in the sidewalls of the furnace. The furnace is smaller thanthe Outokumpu furnace and has only one shaft,which is located at the center of the roof andtakes the off-gas to the cooling system. Usingpure oxygen makes the off-gas volume muchsmaller than in the Outokumpu furnace. Fig-ure 17 shows a sectional view of an Inco flashfurnace. The burner is completely different fromtheOutokumpu type. It has low velocity of about30m/s into the furnace. This gives a short flameand also low flue dust generation. The Inco fur-nace is not equipped with other fossil-fuel burn-ers. Because of the small size of the furnace and

Copper 29

the surrounding equipment, this process is suit-able for replacing old furnaces within existingsmelters.

Operation. Dry concentrate and flux are fedcontinuously to the Inco furnace. The productsof the process arematte (50 – 60%Cu), slag, andoff-gaswith 70 – 80%SO2. The slag is tapped atthe sidewalls of the furnace, and the matte in themiddle part under the offtake. The temperaturesare similar to those of the Outokumpu furnace.In contrast to anOutokumpu furnace the amountof energy which is removed with the off-gas ismuch lower because of the small off-gas quan-tity. Therefore the Inco furnace is not equippedwith awaste-heat boiler for energy recovery. Theoff-gas is cooled by different techniques suchas splash towers, cyclones, or scrubbers [229].The absence of a waste-heat boiler avoids out-of-service times of the furnace. Furthermore, thecopper content in the slag of the Inco furnace islower (0.5 – 1.5 %), and a slag-cleaning furnaceis not needed. Only Hayden Smelter operates anadditional electric furnace. Some operating dataof the Chino, New Mexico furnace are listed inthe following:

Furnace commissioning date 1984Size, inside brick, m hearth w× l× h 5.6× 23× 5gas-off take w× h 3.9× 7.7Concentrate throughput, t/d 1520Blast temperature, C 40Oxygen enrichment, % 98Blast flow-rate, m3/h (STP) 12 500Matte grade, % Cu 58Off-gas volume, m3/h (STP) 20 000SO2 Content 75Recycling of converter slag in flash furnaceSlag treatment discardedFossil fuel input 0

Process Control. The aim of running theInco furnace is the same than the Outokumpufurnace: highest throughput rates with constantfurnace properties such as matte concentrationand furnace temperatures. In comparison withthe Outokumpu process, it is more complicatedto keep the matte grade constant because theoxygen enrichment is not adjustable. This couldbe done by adjusting the mix of the feed (highor low in copper) or adding fossil fuel for oxy-gen consumption. But this is in conflict with theprinciple of the process and does not seem eco-nomic.

5.5.3. KIVCET Cyclone Smelting

Developments in power-plant technology haveled to adoption of the cyclone principle by themetallurgical industry. The acronym KIVCETuses the initial letters of the following Russianterms: oxygen, vortex, cyclone, electrothermic.The development began in 1963, and the firstplant was operated by Irtysh Polymetal Com-bine in Glubokoe, Kazakhstan. The process isnot widely used for copper smelting.

The method is aimed at processing coppersulfide concentrates that contain considerableamounts of other metals. The essential part ofthe continuously operated plant is the smelt-ing cyclone, in which the concentrates are fedvertically, and technical-grade oxygen (95 %)is blown in horizontally, so that reaction takesplace rapidly above 1500 C.

The furnace is divided to allow separating andsettling of the reaction products, in this respectsimilar to a reverb. In contrast to the separa-tion chamber, the atmosphere in the electricallyheated settler ismaintained in a reducing state sothat the slag does not need special posttreatment.

The off-gas volume is small, and the contentof SO2 can be up to ca. 80 %. Metals are reco-vered from the flue dust of both the separatingchamber and the settler. Table 13 and Figure 18explain the process.

Table 13. KIVCET process: analysis and yield [84]

Analysis and yield Metal

Cu, wt % Zn, wt % Pb, wt %

AnalysisConcentrate∗ 25.6 10.0 1.7Copper matte 50.0 2.5 2.0Slag 0.35 3.5 0.2

YieldIn matte 99.1 12.7 60.0In oxidic condensate – 71.0 34.1

∗ Also 24.0 % Fe and 33.0 % S.

5.5.4. Contop Matte Smelting

Using a KIVCET license, KHD HumboldtWedag AG, Cologne developed the combinedContop process [85]. Contop stands for continu-ous smelting and top blowing. It is a combination

30 Copper

Figure 18. KIVCET furnace [84]a) Smelting cyclone; b) Separating chamber; c) Cyclone waste-gas offtake; d) Partition wall; e) Settling reduction hearth;f) Slag tap hole; g) Feed of reductant and offtake for waste gas from the hearth; h) Electrical resistance heating; i) Mattetaphole

of a high-intensity smelting process in a cycloneand secondary treatment of the molten phasewith a top-blown jet. Contop-type cyclone burn-ers have been used on reverberatory furnaces,for example, at Chuquicamata and Palabora. In1993 a complete Contop smelter started opera-tion at El Paso (Asarco) [230], replacing the oldreverberatory furnace. The furnace is shown inFigure 19. It is a kind of a hearth furnace. Twocyclone burners are placed on the top of one endof the furnace. Each is fed with about 25 t/h ofdry concentrate and flux and about 95 % oxygenblast. Natural gas is fed to the cyclone for energyadjustment. The cyclones are made of stainlesssteel and are water-cooled and about 2m in di-ameter and 4m in height. Inside they are pro-tected by a layer of solidified matte – slag. Mattewith 55 – 60 % Cu is produced. At the otherside of the furnace the off-gas is passed througha waste-heat boiler. The amount of off-gas isabout 65 000m3/h (STP) with about 24 % SO2.Its temperature is about 1150 C. After cooling,the gas is washed and sent to sulfuric acid pro-duction.

To keep the slag hot, oxy-fuel burners are in-stalled on the roof of the furnace between thecyclones and the off-gas shaft [231]. The slagcontains 0.8%Cu and can be discarded. The ad-vantage of the Contop cyclone is high flexibilityin concentrate composition, low dust production(ca. 2% of the feedmaterial), high volatilizationof zinc and bismuth from dirty concentrates, and

potential steam generation from the cyclones.Compared with Outokumpu flash smelting, theContop process is more energy intensive (higherspecific cost per tonne of concentrate). Anotherdisadvantage is the relatively short lifetime ofthe cyclone (about one year). Therefore only oneplant was built until now.

Figure 19. Contop furnace

5.5.5. Flame Cyclone Smelting

The flame cyclone reactor (FCR) process wassuggested by LURGI and Deutsche BabcockAG ca. 1975 and has been demonstrated at a pi-lot plant of Norddeutsche Affinerie in Hamburgwith a capacity of ca. 10 t/h. It is a high-tempera-ture (> 1500 C) reaction for autogenous smelt-ing of copper sulfide concentrates in a high-oxygen atmosphere (up to ca. 75 %). A sec-ond characteristic is the simultaneous removal

Copper 31

of volatile compounds of other elements, suchas zinc, cadmium, tin, lead, arsenic, antimony,and bismuth, as oxides and basic salts, in the fluedust. The SO2 content of the off-gas is greaterthan 50 %. The products are a high-grade matte,containing up to 80 %Cu, and slag, which sep-arates in a settler [86].

The principle of this method differs from thatof theKIVCET process in that the reaction in theFCR process takes place in a special chambersituated before the cyclone, where the moltendroplets are separated by centrifugal force. Theprocess is well suited for processing complex ordirty concentrates, but until now never installedin practice.

5.6. Discontinuous Matte Conversion

Matte produced by smelting processes is treatedin themolten state by blowingwith air; this stageof concentration is known as converting. Copperand iron sulfides, the main constituents of matte,are oxidized to a crude copper, ferrous silicateslag, and sulfur dioxide.

The batch converting process has been em-ployed for many decades in two operating stepsat ca. 1200 C in the same vessel. Investigationsand development of continuous methods are be-ing made [87], [88].

The conventional converting of matte isa batch process that yields in the first stepan impure copper(I) sulfide containing ca.75 – 80wt %Cu, the so-called white metal, andin the second step the converter, or blister, cop-per averaging 98 – 99wt%Cu. The name blistercopper derives from the SO2-containing blistersthat are enclosed in the solidified metal.

First Step. The main reactions are oxidationof iron(II) sulfide and slagging of iron(II) oxideby added silica [7631-86-9] flux:

2 FeS + 3O2 −→ 2 FeO+ 2 SO2

2 FeO+SiO2 −→ Fe2SiO4

Formation of magnetite occurs near the tuyeres:

3 FeS + 5O2 −→ Fe3O4 + 3 SO2

Copper(I) sulfide is partially oxidized, but it isalso reformed

Cu2S+ 1.5O2 −→ Cu2O+SO2

Cu2O+FeS −→ Cu2S+FeO

In Figure 6, the first step corresponds to movingalong the pseudobinary Cu2S – FeS line to forman impure copper(I) sulfide.

Second Step. Continuing oxidation occursas in a typical roasting reaction process:

In Figure 6 and Figure 20, the compositionmoves along the Cu2S –Cu line from the cop-per(I) sulfide to crude metallic copper, the twophases being immiscible.

Figure 20. The Cu –Cu2S system [89]

The blister copper contains < 0.1wt %S, ca.0.5wt %O, and traces of other impurities.

Converter Slags. The slags from the firststep are iron(II) silicates (40 – 50wt %Fe) withhigh magnetite content (15 – 30wt %Fe3O4 ).The initial copper content of 3 – 8wt % can in-crease up to 15wt % at the end of the reactionby formation of copper(I) oxide. This slag can

32 Copper

be decopperized by returning it to the smeltingunit or by froth flotation (cf.page 28).

The high-viscosity small-volume converterslags from the second step have a high coppercontent (20 – 40wt %) in the form of copper(I)oxide or silicate.When enough slag has accumu-lated, it is returned to the first converting stage.

Temperature. Converting is a stronglyexothermic process that can overheat duringoxidation of iron-rich mattes. The temperaturemust be held ca. 1200 C by adding fluxes, cop-per scrap, precipitates from hydrometallurgicaltreatment (e.g., cement copper), or concentrates.The off-gas (5 – 10 vol %SO2 ) is transferred toa sulfuric acid plant.

The blowing time per batch is a few hours;however, as the copper content of the matte in-creases, the converting time decreases. Occa-sionally, oxygen-enriched air is used to increasethe throughput.

Impurities. The distribution of other ele-ments among the phases during converting is asfollows:

1) Noble metals and most of the nickel, cobalt,selenium, and tellurium collect in whitemetal and then in blister copper.

2) The bulk of zinc and some nickel and cobaltcollect in converter slag.

3) The oxides and sulfates of arsenic, antimony,bismuth, tin, and the basic sulfates of lead arefound in flue dust.

Converter Types. The copper converter wasinvented in 1880 byManhes and David, basedon the Bessemer converter, which had been usedin the steel industry since 1855. This develop-ment led to the incorrect name “copper besse-merizing,” although the true Bessemer processis a refining step. Originally, the copper con-verter was upright, and such obsolete units werein operation until the early 1980s, e.g., the GreatFalls converter developed by Anaconda MiningCo., United States.

The following types are in use currently(Fig. 21):

1) The Peirce – Smith converter has been themost important apparatus for converting formany decades, and the number in operation

may be in the range of a thousand. Morethan 80 % of the worldwide produced cop-per comes from this type of converter. It isa horizontal cylinder lined with basic bricks(magnesite, chrome –magnesite) that can berotated about its long axis (Fig. 22); blast airis blown through a horizontal row of tuyeres.In practice, the punching of tuyeres with spe-cial devices is necessary to maintain the flowof air. The largest vessels are 12.5m longwith a diameter of 4.6m.

2) Hoboken or syphon converter [91]. This vari-ation of the P – S type was developed yearsago by Metallurgie Hoboken N.V., Belgium,but is now used by only a few smelters in Eu-rope and in North and South America; largerunits are operated at Glogow smelter in Sile-sia, Poland; Cyprus Miami Smelter in Ari-zona; and Paipote smelter (ENAMI), Chile[92]. Its advantage is its freedom from suck-ing in air, so the off-gas can attain SO2 levelsup to 12%. Special features of the design arethe small converter mouth and the syphon orgoose neck that guides the off-gas and fluedust flow.

3) Inspiration converter [93]. Thevessel has twomouths, the smaller for charging, the largerfor the off-gas. The latter is well hooded inall operating positions. It is only in operationat Cyprus Miami Smelter.

4) Top-blown rotary converter [95]. The TBRC,which is known in the steel industry as theKaldo converter, was adopted by the nonfer-rous industry (first by INCO, Canada) be-cause of its great flexibility. Air, oxygen-enriched air, or on occasion, commercialoxygen is blown through a suspended water-cooled lance onto the surface of a chargeof copper-containing materials. In practice,the TBRC is used batchwise for special op-erations on a small scale, but generally notfor converting copper matte. Tests of directsmelting of concentrates (clean, complex,or dirty) to white metal or blister copperwere performed at Ronnskar smelter (Boli-denMetall AB), Sweden, but the processwasnot realized technically. Copper extractionfrom copper scrap and other secondary ma-terials is also carried out. The TBRC is alsowell suited for lead/precious metals metal-lurgy.

Copper 33

Figure 21. The evolution of the copper converter [90]

Figure 22. Schematic cross section and back view of a Peirce – Smith converter [96]

34 Copper

Developments for Increased ProductivityandEnvironmental Protection To increase theproductivity of a Peirce – Smith converter, usu-ally the oxygen enrichment in the blast is in-creased. Depending on process step (slag or cop-per blowing) up to 30%O2 is used. The resultingincreases in the temperature in the vesselmust becontrolled. Copper smelters mostly in Europe orJapan add copper scrap during the copper blowphase to adjust the temperature in the converterbath. Special systems (lift and conveyor) havebeen developed to add up to 70 % of the coolingmaterial without stopping blowing [232], [233](Figure 23).

Figure 23.Machine for automatic charging of cooling ma-terial

Automatic temperature measurement (opti-cal spectrometer or thermocouple through thetuyere) is used for process control and to avoidhigh erosion of the brick lining of the con-verter. Also the chemistry of the convertingprocess is monitored by optical spectrometry[234]. This allows the blowing time to be easilycontrolled without overblowing the charge andthus minimizes the copper loss to the converterslag and magnetite formation in the slag. High-velocity tuyeres have been developed [235],[236] which can be operated with high oxygenenrichment without extensive refractory ero-sion. At the front of the tuyere a tubular accre-tion is formed which protects the brick lining.Another advantage is that these tuyeres need notbe punched (saves labor and maintenance cost).A disadvantage is the relatively high pressure ofabout 3 bar of compressed air (high investment).These high-velocity tuyeres have been tested inPeirce – Smith and Hoboken converters. Today(ca. 2000) the first smelters (Hidalgo) are plan-ing to install the system.

One of the problems of Peirce – Smith con-verters is leakage SO2-containing off-gas duringcharging and pouring operations in the workingenvironment. To avoid these fugitive emissions,special secondary hooding systems have beendeveloped and installed. For example, Figure 24shows the hooding system used at NorddeutscheAffinerie. The collected fugitive gases are dilutein SO2 (up to 0.2 %) and are treated by vari-ous techniques (scrubbing with basic solutionsor dry absorption of the SO2 producing gypsum)[237], [238].

5.7. Continuous Matte Conversion

While continuous copper matte smelting pro-cesses have been in operation for many years,continuous converting of matte has come intouse slowly. The potential benefits are mini-mization of materials handling (especially over-head crane transport of liquids), more efficientoff-gas capture, and continuous SO2 produc-tion for the sulfuric acid plant. Two multiple-furnace processes are used today for continuoussmelting and converting: the Mitsubishi processand the Kennecott/Outokumpu Flash convertingprocess. Noranda has developed a continuouslyrunning converter which has been in operationsince 1997 at Horne Smelter.

5.7.1. Noranda Process

In the 1970s Noranda started with the Norandareactor for directly smelting blister. This wasnot useful and was therefore switched to smelt-ing high-grade matte. A second reactor similarto the smelting one is proposed for continuousconversion ofmatte (patented 1985 [239]). Since1997 it has been operated at Horne smelter. It isa horizontal cylindrical vessel with two mouths(one for adding liquid matte by ladle, the otherfor the off-gas) and a row of submerged tuy-eres. There is also the possibility to feed solidmatte or coal by slinger belt through one endwall. The converter is usually fed with liquidhigh-grade matte (70 % Cu) from the Norandasmelting reactor. It produces semi-blister cop-per with high sulfur content (1 – 1.5 %). Thesemi-blister is poured into a ladle which is trans-ported by special ladle car to the anode furnace.The Noranda converter can be operated in four

Copper 35

Figure 24. Hooding system for the prevention of fugitive emissions at Norddeutsche Affinerie

modes: with molten matte feed, with any com-bination of solid and molten matte feed, up to100% solid feed, in smeltingmodewith concen-trate like the Noranda reactor and a conventionalPeirce – Smith converter. The converter at HorneSmelter (Figure 25) [240] is 4.5m in diameterand 19.8m long (inside the brick lining). It has42 tuyeres. The process off-gas is collected in awater-cooled hood and sent to the sulfuric acidplant. The converter has a secondary ventilationsystem, the gases of which are sent directly toa stack. This process operates continuously andhas very high flexibility but not all of the poten-tial benefits have been achieved. There is stillcrane transport and handling with ladles whichcauses fugitive emissions.

5.7.2. Mitsubishi Process

Mitsubishi Metals Corp., Japan, tested thenew concept during the 1960s and started thefirst commercial plant at Naoshima smelter in1974. Today (ca. 2000) four companies are op-erating this process (Naoshima, Japan; KiddCreek, Canada; LG Metals, Korea; Gresik, In-donesia). The principle of this process is the in-terconnection of three furnaces, as shown in Fig-ure 26 [241], [242]:

1) Smelting or S-furnace: Dried concentrates,flux material, pulverized coal, return con-

verter slag, copper scrap, and flue dust aresmelted in this furnace. The fine material(concentrate, coal) is fed through nine orten vertical steel lances on the top of themolten bath. The blowing lances consistof two concentric pipes. Through the innerpipe, the dried concentrate is air blown, andthrough the outer one, oxygen-enriched blast(40 – 50%O2). The lances are rotated to pre-vent them sticking in the roof. They extendto 0.5 – 1m above the molten bath. About0.3 – 0.5mof the lance is consumedeachday.The off-gas contains 30 – 45 % SO2 and issent to sulfuric acid production.

2) Slag-cleaning furnace: Slag and matte fromthe S-furnace flow continuously by grav-ity into the elliptical slag-cleaning furnace,which is an electric furnace with three orsix submerged graphite electrodes. The tem-perature of the slag is kept at 1250 C.The slag is decopperized to 0.6 – 0.9 % Cuand discarded. The matte flows continuouslythrough a siphon into the converting furnace.

3) Converting or C-furnace: The matte (68 %Cu) is converted continuously by blowingenriched air (30 – 35 % O2) and CaCO3 fluxthrough six lances on the top of the bath. Alsocopper scrap is added through the sidewallof the furnace. Conversion only takes placewhere the oxygen comes into contactwith the

36 Copper

Figure 25. Noranda converter at Horne Smelter

sulfides. This is achieved by running the con-verterwith a special basic calcium ferrite slag(40 – 50 % Fe, 15 – 20 % CaO). The usualsilica slag is not possible because when oxy-gen is blown onto the top, solid magnetite isformed and blocks the surface. In the calciumferrite slag dissolves the magnetite but also alarge amount of copper (15 – 18%). The con-verter slag is returned to the S-furnace. Theblister copper contains slightly more sulfur(0.7 %) than from a Peirce – Smith converter(0.02 %). It is sent to an anode furnace.

Some operating data for the Mitsubishi pro-cess (Naoshima Smelter, Japan) are summarizedin the following:

Commisioning date 1991

S-furnaceDiameter inside brick, m 10Number of lances 9Concentrate throughput, t/d 2050Copper scrap, t/d 20Blast, m3/h (STP) 37 500O2 enrichment, % 49Off-gas volume, m3/h (STP) 39 000SO2 in off-gas, % 29Matte grade, % 68

Slag-cleaning furnaceSize, inside brick w× l× h, m 6× 12.5× 2Number of electrodes 6Residence time ca. 2 hSlag 0.6 % CuOff-gas volume, m3/h (STP) 50

C-furnaceDiameter inside brick, m 8Number of lances 10oxygen enrichment, % 33Blast flow rate, m3/h (STP) 25 500Copper Scrap, t/d 190Off-gas volume, m3/h (STP) 25 000SO2 in off-gas, % 26

The major advantage of the Mitsubishi pro-cess are the good SO2 capture and the lowerhandling expense of materials. A disadvantageduring the initial operating time was that noscrap material could be added. This is has nowbeen solved [243]. With a smelting capacity ofabout 2000 t/d of copper concentrate, NaoshimaSmelter processes also about 40 000 t/a of cop-per scrap from the market and additional anodescrap from the refinery. Another problem is theimpurity behavior. Especially lead is a problembecause of the calcium ferrite slag. If too muchlead is in the feed, the copper anodes are too richin lead for electrorefining. A comparison of theimpurity behavior is given in Table 14.

Table 14. Impurity contributions to anodes (wt % of input)

Process Reverbatory/Peirce – Smith Mitsubishi

Pb 8 – 13 15 – 19As 9 – 11 4 – 6Sb 28 – 34 15 – 20Bi 14 – 26 15 – 25Ni 45 – 50 67 – 75

Copper 37

Figure 26.Mitsubishi process

5.7.3. Kennecott/Outokumpu FlashConverting Process

In the 1980s Kennecott announced the develop-ment of their Solid Matte Converting (SMOC)process [244]. This process is based on solidify-ing the molten matte from the smelter, grinding,and feeding the solid to an high-oxygen blastconverter. This process decouples the smeltingand converting steps and gives great flexibility.Later on, it was developedwithOutokumpu, andfor the converting step also a well-proven flashfurnace was chosen. In 1995 the process wentinto operation at Garfield Smelter.

Formatte smelting a flash furnacewith a con-centrate feed rate of about 140 t/hwas built (oxy-gen enrichment 65 – 75 %). The liquid mattecontains 68 – 70 % copper and is granulated inwater. In the second step the matte is milled andthen fed together with lime as flux material tothe flash converter, which operates with an oxy-gen enrichment of 65 – 75 % O2. The feed rateof matte is about 60 t/h (up to max. 80 t/h) [245].This furnace is constructed like the smelting fur-nace. A blister copper and a calcium ferrite slagwith about 18 % copper (like in the Mitsubishiprocess) is obtained. The slag is granulated andfed back to the flash smelter.

The off-gases contain > 30 % SO2 and areconverted to sulfuric acid (input concentration of14%SO2 in the sulfuric acid plant). This processhas very low SO2 emissions of only 3.5 kg SO2per tonne of copper, which is until now the low-

est in the world. This is achieved because it hasalmost no fugitive emissions. The disadvantageis that no scrap can be added to the converter.

5.8. Direct Blister Smelting

Normally copper extraction takes place in twosteps – concentrate smelting and matte convert-ing – in which the chemical reactions are prin-cipally the same: oxidation of Fe and S. It haslong been the aim to combine these processesin a single step/reactor and minimize the en-ergy consumption and operating cost. Today (ca.2000) direct blister smelting is carried out at twosmelters (Glogow, Poland and Olympic Dam,Australia) in a Outokumpu flash furnace. Be-tween 1973 and 1975, Noranda also used theirreactor for blister smelting from concentrate, butit was switched to matte smelting because ofimpurity problems. Another proposal is the QSprocess.

5.8.1. Blister Flash Smelting [246]

This process takes place in a conventional Out-okumpu flash smelting furnace. Concentrate andsilica flux are fed with 60 – 85 % O2 blast tothe burner. The amount of O2 is just enough toform blister copper. In practise, the particles inthe reaction shaft are somewhat over-oxidizedon the outside and incompletely oxidized on theinside. In the molten bath layers the reaction is

38 Copper

completed to form blister copper and iron silicaslag. Formation of a layer of matte between blis-ter and slag must be avoided, because otherwisea pronounced foaming reaction could take place.

In both operations copper concentrates withlow iron content (e.g., chalcocite, bornite) areprocessed to minimize slag production and cop-per loss in the slag. The slag has a high cop-per content (15 – 21 % Cu), mainly in oxidizedform, and has to be reduced in an electric fur-nace (produces copper – iron – lead alloy [247])or concentrated by grinding and flotation. Re-duction in an electric furnace with addition ofcoke takes about 12 h residence time. The fur-nace is operated at higher temperatures than thenormal flash smelter (slag temperature 1300 C)to obtain higher solubility of magnetite in theslag and decrease the viscosity of the slag. Theblister copper contains 0.3 – 1 % S. At the Glo-gow Smelter, the process has been in operationsince 1977 with a daily throughput of 1500 t ofconcentrate. At Olympic Dam it has been oper-ated since 1988 and was enlarged in 1998.

5.8.2. QS Process

The QS process, based on the QSL leadmakingprocess, was invented by Queneau and Schuh-mann in 1974 [109]. Blister copper is producedfrom concentrate in a single vessel. The reactoris similar to the Noranda reactor but has counter-current flowof slag andmatte.Gases are injectedfrom the bottom [248] of the vessel to oxidizeFe and S and to reduce the copper content in theslag (Figure 27). Although the process has fun-damental and technical merits, it has not beenrealized in a commercial plant.

5.9. Copper Recycling

Most copper (> 95 %) is used in metallic form,as copper metal or copper alloys. Recycling ofcopper and its alloys has been carried out sinceancient times. In 1997, 37%of copper consump-tion came from recycled copper. This low figurecould be explained by the long lifetime of cop-per products (e.g., more than 30 years for copperwire) and the much increased copper productionin the last 50 years. Virtually all products madefrom copper can be recycled, and copper can be

recycled over and over again without losing itschemical or physical properties.

The processes used for copper recycling de-pend on the copper content of the secondaryrawmaterial, its size distribution, and other con-stituents (Table 15) [249], [250]. Three generaltypes can be defined:

–Type 1: copper scrap (new and old scrap), usedfor smelting and refining or direct smeltingfor products. This type accounts for about95 % of all recycled copper. The value of themetal is much higher than its treatment costs.

–Type 2: copper-containing special scrap such ascables and printed circuit boards. Pretreat-ment (e.g., cable comminution) is necessarybefore smelting the copper. Another exampleis copper in cars. Each car contains 10 – 30 kgof Cu which must be separated from the iron.The value of the material is in the range ofthe overall treatment cost.

–Type 3: copper-containing residues, for exam-ple sludges from metal-plating industry. Thecopper content of these materials is low. Thevalue of the copper is much lower than thetreatment cost of the material.

High-purity secondary copper material isused for smelting and casting new products, andno fire and/or electrolytic refining is necessary.This is also true of pure alloy scrap, which isused for fresh alloy. With increasing impuritycontent, the material is fed to smelting and refin-ing processes. The process steps and aggregatesof secondary smelters are generally similar tothose of primary production, but the secondaryraw material is usually metallic or oxidic [253].Therefore, mainly reducing conditions are usedfor secondary smelting.Often carbon in the formof coke or natural gas is used.

For smelting metallic scrap, often shaft fur-naces (Asarco type or Contimelt) are used. Forsmelting oxidic materials, blast furnaces, elec-tric furnaces, TBRCs, and Isasmelt furnaces areused. In the TBRC both the smelting and theconverting step could be done. This saves en-ergy in comparison with the blast furnace andPeirce – Smith converter. In some cases electricarc furnaces are used instead of the blast fur-nace. The advantage is that this furnace producesabout 80 % less off-gas than the blast furnace.Thismakes the capture easier andmore efficient.

Copper 39

Figure 27. QS process

Table 15. Overview of different sorts of secondary copper sources

Type Material Cu content, % Source Recycling process

1 mixed copper scrap 90 – 95 sheets, gutters, water boilers, heaters,wires, pipes

converting, fire refining

1 pure copper scrap 99 semi-finished products, wire, strip,cuttings

melting and casting of semifinishedproducts

1 red brass scrap 75 – 85 valves, taps, machine components,fittings, bearing boxes, car radiators

converting or alloy production

2 shredder material 60 – 65 cars blast furnace, electric furnace, TBRC2 cable 40 – 50 buildings, cars shredder, firerefining2 electronic scrap 10 – 20 electronics converting, TBRC3 sludges 5 – 10 electroplating blast furnace, Electric furnace3 copper – iron material 10 – 20 armatures, stators, rotors blast furnace, electric furnace, TBRC3 drosses, ashes, slags 20 – 25 foundries blast furnace, electric furnace, TBRC

A typical flow sheet of a secondary smelter isshown in Figure 28.

Many primary smelters treat scrap besidestheir internal recycling materials like anodescrap from the tankhouse or converter flue dusts.Additional copper scrap bundles are fed to thePeirce – Smith matte converter. The scrap coolsthe converters and uses the energy from matteoxidation for heating and melting.

Also electronic scrap such as printed circuitboards is fed to copper matte converters (e.g.,Norddeutsche Affinerie [251], [252]) or No-randa reactors (e.g., Noranda). The copper andthe precious metals are captured and sent to sub-sequent refining. The organic polymers are burntat the high temperatures, and water and CO2 areformed; ceramics are slagged. Because of thehigh temperature and the high SO2 content inthe off-gas, no dioxins are formed.

5.10. Hydrometallurgical Extraction

About 15 % of the primary copper productionin the western world, that is, about 1.8× 106 t/a,

results from sulfuric acid leaching of copper ore,which is combined with solvent extraction andelectrowinning. Other proposed techniques likeammonia leaching or hydrochloric acid leachinghave minor importance. They are mainly pro-posed for treating concentrates. A typical flowsheet of a sulfuric acid leaching operation isshown in Figure 29.

Chemistry Of Sulfuric Acid LeachingProcesses. There are two chemically differentclasses of copper minerals:

1) Oxidic or secondary minerals like malachiteand chrysocolla

2) Sulfidic or primary minerals like chalcopy-rite and chalcocite

The sulfidicminerals are themost common in anore body. Oxidic minerals are often located onthe top of the ore body and are present in smalleramounts.

Chemical reactions during extraction of cop-per from primary and secondary minerals are asfollows:

40 Copper

Figure 28. Flow sheet of a secondary smelter

Figure 29. Sulfuric acid leaching of copper ores

Copper 41

Oxidic secondary copper minerals:

CuCO3 · Cu(OH)2 + 2H2SO4 −→2CuSO4 + 3H2O+CO2

Sulfidic primary copper minerals:First step: generation of Fe3+, which is the

oxidant for copper leaching, by bacteria.Thiobacillus thiooxidans:

2 FeS2 + 7O2 + 2H2O −→ 2 FeSO4 + 2H2SO4

Thiobacillus ferrooxidans:

2 FeSO4 + 0.5O2 +H2SO4 −→ 2 Fe3+ + 3 SO2−4 +H2O

Second step: attacking the chalcopyrite:

2 Fe2(SO4)3 +CuFeS2 + 3O2 +H2O −→5 FeSO4 +CuSO4 + 2H2SO4

Secondary minerals are readily soluble in sulfu-ric acid, and reaction times are typically short.After several hours most of the copper is ex-tracted, but sometimes the ore requires a fewweeks reaction time. Recovery of copper fromprimary minerals is more difficult because thesulfides are very stable, and apart from sul-furic acid an oxidant is required as well. Themost important oxidant is Fe3+, which is gener-ated from the mineral chalcopyrite by bacteria-assisted reactions with atmospheric oxygen. Al-though these bacteria, which normally exist inthe mine water, speed up the process, very longreaction times like months or years are still nec-essary to extract most of the copper. The con-sumption of sulfuric acid depends on the type ofcopper mineral. The bacterial leaching processproduces sulfuric acid itself.

Industrial Leaching Processes. The copperleaching techniques in practical use are listed inTable 16.

The most important technique is heap leach-ing. It is very common for oxidic copper ore. Thetypical copper content is between 0.25 and 1 %.The ore is crushed to 1 – 10 cm and stacked toheaps 3 – 10m high with an area of about 10 000to 100 000m2. Dilute sulfuric acid is sprayed onthe top of the heap. The liquor trickles throughthematerial, copper is dissolved, and the copper-bearing solution (1 – 5 g/L Cu) is collected at thebottom of the heap. This copper solution is puri-fied and enriched by solvent extraction. Copper

metal is generated by electrolysis of the enrichedsolution. After several months (1 – 6), leachingis stopped and a new heap is placed on the topof the previous and leaching is begun again.

Data from a typical heap leaching plant areshown in Table 17. To produce 54 000 t/a of cop-per, 27 000 t ore per day must be treated.

Table 17. Typical data of heap and dump leaching operation

Heap leaching Dump leaching

Cu cathode production, t/a 54 000 6600Leach ore composition chrysocolla 95 %

chalcopyrite, 3 %chalcocite

Soluble Cu in ore, % 0.54 0.15Total area under leach, m2 200 000 300 000Tonnes of ore per day 27 000 55 000Annual Cu recovery rate, % 90 20Sulfuric acid consumption,t/t Cu

1.7 0

The copper production of a typical dump-leaching operation is much lower, for example.6600 t/a. The feedmaterial is very low grade sul-fidic ore, and several years are required to extractall the copper. The sulfuric acid consumptiondepends on the mineralization of the copper oreand the composition of the host rock. Typicalfigures are in the range of 1 – 5 t of sulfuric acidper tonne of copper product.

Minor methods are in situ leaching, tailingsleaching, and vat leaching. Vat leaching wascarried out in the past to produce a pregnantleach solution with about 30 g/L Cu which wasdirectly suitable for electrowinning. This tech-nique is replaced by heap leaching and SX. Agi-tation leaching is usedmainly to leach oxidic Cuconcentrates in Zaire and Zambia. Fine particlesare treated with strong sulfuric acid (ca. 60 g/L)solution, and the reaction is complete after sev-eral hours.

Other Leaching Processes. Mainly for theleaching of sulfidic copper concentrates, manyprocesses have been developed.

The Arbiter/Escondida Process was devel-oped for leaching mainly chalcocite concen-trates. In ammonia solution, half of the copperis extracted very quickly and CuS remains in theresidue.

Cu2S+ 2NH3 + 2NH+4 + 0.5O2 −→

CuS+ [Cu(NH3)4]2+ +H2O

42 Copper

Table 16. Common techniques for sulfuric acid leaching of copper mine

Leaching technique Mineralization % Cu in ore Leaching time Estimated Cu production(1998), T/A

Heap oxides, chalcocite 0.25 – 1 several months to years 700 000Dump chalcopyrite < 0.25 1 – 5 decades 50 000In Situ all > 0.5 decades 50 000Vat oxides 1 – 2 5 – 10 d 50 000Agitation oxides (carbonates) 20 – 40 2 – 5 h 200 000Tailings oxides (carbonates) 0.25 – 1 1 d 100 000

The reaction takes place in agitated vessels. Thecopper complex solution is sent to a solvent ex-traction and electrowinning plant. The residueis sent to a flotation plant, where copper is en-riched, and then fed to a smelter. This processwas operated at Escondida mine in Chile but atpresent is shut down.

The Cuprex process uses iron(III) chloridefor the oxidation of copper from chalcopyrite:

CuFeS2 + 4 FeCl3 −→ CuCl2 + 5 FeCl2 + 2 S

A solvent extraction step gives pure CuCl2 so-lution (ca. 100 g/L Cu), which is fed to a mem-brane-divided electrowinning cell, where cop-per powder is produced and in at the anode theiron(II) chloride is oxidized to iron(III) chlorideand recycled to the leaching process. The pro-cess has been tested in a 1 t/d pilot plant.

The Intec process is carried out in aNaCl/NaBr solution at 80 – 85 C with CuCl2as oxidant. Leaching is carried out in four steps;gold is also leached in the last step. The residueconsists of Fe(O)OH and elemental sulfur.

4CuFeS2 + 4CuCl2 + 3O2 + 2H2O −→8CuCl + 4 FeOOH+8S

The solution is cleaned by adding CaO and thensent to a membrane electrolytic cell. At the cath-ode copper is reduced to copper powder. At theanode two reactions takeplace: oxidationofCu+

toCu2+ and oxidation of bromide ions toBrCl−.This solution, which contains about 30 g/L Cu,is sent back to the leaching reactor (step 4, Auleaching).

Oxygen pressure leaching of chalcopyrite,bornite, or chalcocite concentrates in sulfuricacid under 8 bar oxygen pressure at 200 Cgivesabout 99 % copper recovery after a short reac-tion time of only one hour. The dissolved coppercan be recovered by electrowinning. The leach

residue can be treated with cyanide for gold re-covery. This process was tested in laboratoryscale. It might find use for impure concentratesor concentrates high in precious metals.

Bacterial leaching of chalcopyrite concen-trates (Bio COP) is developed since 1997 atCodelco’s Chuquicamata mine. The sulfide con-centrate is leached in stirred tanks at 65 – 85 Cassisted by special heat-resistant bacteria (bac-teriamesofila). The process needs large amountsof air/oxygen. The copper sulfate solution is fedto a SX-plant and then to copper electrowinning.A pilot plant will be built within the next 2 years.

Solvent Extraction. The pregnant leach so-lutions from heap, dump, or in situ leachingare too dilute in Cu (1 – 5 g/L) and too impurefor direct production of high-grade copper cath-odes. Industrial copper winning requires elec-trolytes with about 40 – 50 g/L copper. Copperenrichment and purification, mainly to removeiron, is done by solvent extraction. The dilute,impure solution from the leaching operation ismixed with an organic solution, which is basedon kerosene or petroleum and contains about5 – 10 % of an extractant which is selective forcopper. The organic extractants used for the pro-cess are salicylaldoximes and ketoximes (RH)and commercially available as LIX (Henkel) orMOC (Allied Signal). The copper ions are com-plexed by these reagents and enter the organicphase:

2RH+CuSO4 −→ R2Cu+H2SO4

After mixing, the aqueous and organic phasesare separated by gravity (settler). In a secondstep, copper is stripped from the organic phasewith more highly concentrated sulfuric acid so-lution.

H2SO4 +R2Cu −→ 2RH+CuSO4

Copper 43

Normally, the depleted solution from elec-trowinning, which contains about 170 g/LH2SO4 and 35 g/l Cu, is used. It is enrichedwith copper to about 50 g/L and sent to the elec-trowinning cells. Depending on the compositionof the dilute solution from the leaching plant sev-eral mixer/settler steps are needed for purifica-tion of the copper.

Copper Electrowinning. The copper-enriched sulfuric acid solution (ca. 40 – 50 g/LCu, 140 – 170 g/L H2SO4) is sent to the elec-trowinning cells, which are similar in construc-tion to refining copper cells. Copper is depletedat the cathode on stainless steel sheet (ISA Pro-cess) or copper starters. The cathode reaction isidentical to the refining reaction:

Cu2+ + 2 e− −→ Cu E0 = 0.34V

The anode reaction is completely different. In-ert anodes, mainly PbSnCa alloy, are used, andoxygen is formed at the anode.

H2O −→ 0.5O2 + 2 H+ + 2 e− E0 = 1.23V

The electrical potential needed for copperelectrowinning is about 2V, as opposed toabout 0.3V for copper electrorefining. About2000 kWh is required for depleting 1 t of cop-per. The usual plating period is about 7 d, af-ter which around one-third of the cathodes aretaken out of the cell by crane and immediatelyreplaced by fresh ones, in contrast to the refiningtankhouse, where all cathodes are removed. Thecathode replacement in the winning tankhouseis performed without cutting off the electricalpower. This maintains a passivation layer onthe lead anode and minimizes contamination ofthe copper with lead. In addition, the electrolytecontains about 100 ppm cobalt to prevent corro-sion of the anodes. The lead anodes have life-times of up to 20 years. The current density inelectrowinning plants is about 200 – 300A/m2.Modern electrowinning tankhouses have capac-ities up to 200 000 t/a of cathode copper.

Economics of Copper Production byL/SX/EW Processes Compared with copperproduction in smelters, L/SX/EW processeshave lower production figures. But the argu-ments for the L/SX/EW techniques are low in-vestment costs and increased copper output ef-ficiency of mines. Furthermore, copper can be

recovered from low-grade materials which areunsuitable for smelters. With improvements insolvent extraction and electrowinning technol-ogy, the quality of the recovered copper has be-come as good as that from classical refineries.Leaching facilities are mostly located at the cop-per mines. The disadvantage of leaching is thatit does not recover the precious metals.

Production figures (in 103 t/a) for L/SX/EWcopper by region in 1997 follow:

North America 567Latin America 943Oceania 49Africa 65Asia 4Western Europe 2

The expected growthofwesternworld copperproduction and the contributions from smeltersand L/SX/EW operations is shown in Figure 30.Copper production is expected to increase fromabout 12× 106 t/a in 1996 to about 14× 106 t/ain the year 2000. The biggest part of this growthwill come from L/SX/EW operations. Smelterproduction will also increase, but there the nextstep in capacity enlargement is expected after2002.

Figure 30.World copper production statistics

Most of the increase in copper production byleaching was accounted for by Latin America(Figure 31). By 2002 or 2003, the production ofL/SX/EW copper is expected to have increasedby nearly a factor of three. Several new opera-tions are scheduled for the near future.

44 Copper

Figure 31. Copper production by the L/SX/EW technique

6. Refining

Conventional refining comprises three stages:(1) pyrometallurgical or fire refining, (2) elec-trolytic refining, and (3) remelting of cathodesand casting of shapes. Refining without elec-trolysis is adequate if the fire-refined copperhas the necessary purity and if the content ofprecious metals can be neglected. If extremelyhigh-purity copper is needed, zone melting orrepeated electrolysis of cathodes is used.

6.1. Pyrometallurgical Refining

Fire refining is applied to crude copper suchas blister copper from converters (ca. 97 –99wt % Cu), black copper from blast furnaces(ca. 90 – 95wt % Cu), cement copper from hy-drometallurgical operations (ca. 85 – 90wt %Cu), anode scrap from electrolytic refining, andhigh-grade copper scrap, chiefly unalloyed wirescrap.

The refining of molten copper to anode cop-per for electrolysis or commercial fire-refinedcopper has the following functions:

1) Removing impurities by slagging andvolatilization, with the precious metals re-maining entirely in the metallic copper

2) Reducing the sulfur content to ca.0.0005 – 0.005wt % by oxidation

3) Decreasing the oxygen content to< 0.1 – 0.25 % by reduction (poling) to givea flat surface as a result of the water-gasequilibrium in molten copper

6.1.1. Discontinuous Fire Refining

Two furnace types are available for batch copperrefining, the older reverberatory furnace and themore modern rotary furnace. The former, whichresembles smaller reverbs for matte smeltingfrom concentrates, has a capacity of 200 – 400 tof copper per charge, can be fed with molten orsolid copper, and is used in secondary smelters.Rotary furnaces hold up to 350 t of molten metalper charge and are generally fed only with liq-uid copper. This type of furnace is preferred byprimary smelters because the reduction is moreefficient. An extra melting furnace, e.g., anodeshaft type (Section 6.1.2), can be required forremelting of solid materials (scrap and anoderests).

Low-sulfur pulverized coal, fuel oil, natu-ral gas, or reformed natural gas serve as fuel.The refractory lining consists of basic bricks,such as magnesite or the more spall-resistantchrome –magnesite bricks.

After charging and possibly melting, oxida-tion and reduction stages are carried out in se-quence. At the beginning of the oxidation pe-riod, air is blown into the melt, partly slaggingandpartly volatilizing the impurities.During thisblowing step, a part of the copper is oxidized tocopper (I) oxide, which dissolves in the liquidmetal (Fig. 32). If the content of Cu2O in cop-per increases to ca. 10wt % (corresponding to1wt%O), it acts as a selective oxidizing agent:

Cu2S+ 2Cu2O −→ 6Cu + SO2

In practice, large amounts of SO2 are generatedso that this final stage of the oxidation period istermed boiling. Reduction of the sulfur contentlimits SO2 blisters in the solid copper.

The subsequent poling with wood is acenturies-old method that is still employed inolder reverb plants. Large tree trunks (poles) ofbeech, birch, eucalyptus, etc. are plunged un-der the surface of the melt to generate reducinggases and steam by dry distillation of the wood.The escaping gas mixture reacts with copper (I)

Copper 45

oxide and mixes the molten bath [1, p. 441], [2,p. 392].

This awkward operation has been largely dis-placed by gas poling [143] in rotary furnaces:natural gas (CH4 ), reformed gas (CO, H2, andN2), propane, or ammonia [144] is blown intothe copper melt through tuyeres. This processhas been introduced at most smelters in theworld.

Figure 32. Partial phase diagram of the Cu –Cu2O system

The poling operation proceeds in two steps.During tight poling, the remaining sulfur dioxidefrom copper (I) sulfide is almost entirely flushedout by the escaping gas, a sample of liquid cop-per at the end of this stage solidifying withoutblisters or cavities. Next comes the poling toughpitch,which is necessary to reduce the copper (I)oxide and achieve the required low oxygen con-tent. High Cu2O content in the solidified metalcauses brittleness and decreased strength; more-over, Cu2O disproportionates to copper metaland copper(II) ions in sulfuric acid electrolytes,which disturbs the electrolytic refining opera-tion.

The final oxygen content in fire-refined toughpitch copper must be 0.02 – 0.05wt %; for an-odes it can be 0.05 – 0.3wt %. The fire-refiningprocess can be understood from the Cu –O sys-tem [145] (see Fig. 32); the system Cu –O–S isimportant in the oxidation period (see Fig. 5) and

theCu –H–O system for the reduction or polingperiod.

When a copper melt solidifies, a shrinkageof ca. 5 vol % occurs, but a flat surface can beachieved by careful control of the equilibrium

Cu2O+H2 2 Cu +H2O

The steam of micropores can compensate thevolume difference, and a flat set cast is obtained.

Surface and fracture of small samples of so-lidified copper from the molten bath are ob-served during the refining process to ascertainthe current state. It is also possible to measurethe oxygen content of the copper melt potentio-metrically.

A special problem is the extremely highcopper content (up to ca. 40wt % Cu, chieflyas Cu2O) in refining slags. Such products aretreated as high-grade oxidized copper ores.

6.1.2. Continuous Fire Refining

The two-stage Contimelt process for coppermelting and refining was developed in 1968 byNorddeutsche Affinerie at Hamburg, in coop-eration with Metallurgie Hoboken-Overpelt inOlen, Belgium [147]. The first stage began op-eration in 1979, and the complete process hasbeen operated since 1982 on a commercial scale.The continuous operations are performed suc-cessively in two units connected by launders.First is the anode shaft furnace, where charg-ing, melting, and oxidation of crude copper takeplace. Second is the small drum-type furnace,where poling and casting of anodes are carriedout. The oxygen content of copper from the an-ode shaft furnace averages 0.6 %, with 0.15 %after poling. A feature of the shaft furnace is theadditional equipment with oxygen burners forregulating the composition of the furnace atmo-sphere and the overheating of copper. In com-parison with conventional fire refining, the Con-timelt process provides savings in energy andlabor.

6.1.3. Casting of Anodes

The conventionalmethod of producing anodes isthe discontinuous casting on casting wheel ma-chines. The pure copper molds must be sprayed

46 Copper

with a slurry that prevents the sticking of so-lidified anodes; baryte, alumina, or silica flourare suitable. (Calcium-containing material isnot suitable because it forms gypsum, which ispartially soluble in the electrolyte.) The cast-ing rate can reach 100 t/h. The anode weightsvary between 250 and 450 kg, depending onthe refinery. Anodes from modern plants usu-ally have the following dimensions: 0.9 – 1.1mlong; 0.9 – 1.0m wide; and 3.5 – 6.0 cm thick.They weigh 300 – 450 kg. Economic considera-tions call for anodes of nearly the same weightwithin a plant; therefore, discontinuous castingis best controlled by electronic systems.

Contilanod Process. The Contilanod pro-cess, developed by Metallurgie Hoboken-Overpelt in Olen, Belgium, produces uniformanodes by using the continuous Hazelett twin-belt casting system [148]. The continuous caststrip of anode copper formed between two beltsand damblock chains is 1m wide and 1.5 – 6 cmthick; special cutting equipment separates thestrip into individual anodes 1m long. Some re-fining plants use this method, e.g., White Pine,Michigan, United States; Kidd Creek, Timmins,Canada. The advantage of this system is the uni-formity of the anodes and the high degree ofautomation. But in comparison to the mold-on-wheel technique, the Hazelett caster has some-what higher maintenance costs.

6.2. Electrolytic Refining

About 80 % of the world copper production isrefined by electrolysis. This includes all primarycopper and much secondary copper. This treat-ment yields copper with high electrical conduc-tivity and provides for separation of valuable im-purities, especially the precious metals.

The basic patent, GB 2838, for galvanic de-position of metals was awarded to J. B. Elk-ington in 1865. The most important technicalproblems were solved by E.Wohlwill at theNorddeutsche Affinerie in Hamburg, Germany,in 1876, and this method has been used eversince. The first electrolytic copper refinery in theUnitedStateswasoperated from1883 to1918bytheBalbachSmelting andRefiningCo.,Newark,New Jersey.

6.2.1. Principles [149], [150]

Several possible half-reactions can occur at theelectrodes.

Anode reactions Cathode reactions Standard electrodepotentialE (25 C), V

Cu−→ Cu2+ + 2 e−

Cu2+ + 2 e−

−→ Cu0.337

Cu−→ Cu+ + e−

Cu+ + e−

−→ Cu0.521

Cu+ −→ Cu2+ + e−Cu2+ + e−

−→ Cu+0.153

Secondary reactions occur in the electrolyte:

2Cu+ −→ Cu2+ +Cu (disproportionation)

2Cu+ + 2H+ + 0.5O2 −→ 2Cu2+ +H2O (air oxidation)

Cu2O+2H+ −→ 2Cu+ +H2O (dissolution of Cu2O)

Oxidation by air and disproportionation of cop-per (I) ions yield a surplus of copper (II) ions inthe electrolyte. The coppermetal powder formedby the disproportionation of Cu+ contributes tothe accumulation of the anode slime.

The electrochemical equivalent of copper de-pends on the oxidation state of the copper:

Species gA−1 h−1 mg/L

Cu2+ 1.185 0.3294Cu+ 2.371 0.6588

The greater electrochemical equivalent ofcopper (I) suggests the use of solutions of cop-per (I) ions instead of copper (II) ions. However,this concept has not been put into practice be-cause of enormous industrial difficulties [151].

The two most important electrical parame-ters in electrolytic copper refining operations arethe cell voltage and the current density. The cellvoltage, which usually ranges between 0.25 and0.3V, is determined by several factors:

1) Ohmic resistance of the electrolyte, depend-ing on composition, temperature, electrodedistance, and cell construction

2) Polarization, especially concentration polar-ization of electrodes, which depends on therate of electrolyte circulation

Copper 47

3) Overpotential because of organic additives(e.g., an inhibitor for achieving uniformelec-trocrystallization of copper)

4) Voltage loss in the circuit5) Anode passivity, which may occur at high

current densities

The interaction of these effects is difficult topredict. At any particular electrolytic facility, acontinuing effort ismade to optimize parametersthat affect the cell voltage. The voltage loss in theconductors and contacts is minimized by goodplant design and use of special contacts (Bal-timore grooves, Whitehead contacts, wet con-tacts, etc.).

The second important parameter is the ca-thodic current density. With increasing currentdensity the production of copper increases andthe current efficiency decreases because thecathode potential depends on the current den-sity.

Impurities. The behavior of impurities de-pends on their position in the electrochemicalseries: elements more noble than copper are in-soluble, while less noble ones dissolve in or re-act with the electrolyte. For that reason, the an-odematerial is distributed by electrolysis amongthree phases: cathode copper, electrolyte, andanode slime (Table 18) [152]. Table 19 lists thefractions of anode elements that enter residuesand electrolyte.

Copper Cathodes. Cathode copper is pro-duced currently in a purity between 99.97 and99.99 %. Silver can be deposited in traces; how-ever, this can be avoided by precipitating thesilver from the electrolyte with chloride ion.Other impurities, such as suspended slime ordroplets of the electrolyte, may be mechanicallyoccluded. The following measures are taken toproduce copper of high purity:

1) Maintenance of the optimum current density,to prevent cathodic deposition of other ele-ments (e.g., arsenic)

2) Addition of organic inhibitors to avoid theformation of nodules on the cathode surface

3) Removal of impurities such as arsenic, anti-mony, and bismuth from the electrolyte byadsorption or chemisorption

4) Prevention or elimination of suspendedslimes by regulating the electrolyte flow andoccasionally filtering it.

Electrolyte. Thecompositionof copper elec-trolytes from various plants is generally simi-lar: ca. 35 – 45 g of copper and 150 – 220 g ofsulfuric acid per liter at an operating tempera-ture of 55 – 65 C (see Table 14). As a result ofsecondary reactions during electrolysis, the con-centration of copper (II) ions increases slowly;therefore, this copper surplus must be recoveredby cathodic deposition in a few (ca. 2 %) liber-ator cells equipped with insoluble anodes, usu-ally of antimonial lead. Soluble impurities, suchas iron, cobalt, zinc, manganese, most of thenickel, and some arsenic and antimony, are alsoenriched in the electrolyte.

The upper limits of impurity content are ca.10 g/L for arsenic and 20 – 25 g/L for nickel.Part of the electrolyte is withdrawn continuouslyfrom the circuit for purification, and the volumeis compensated by adding sulfuric acid and cath-ode wash water. There are two methods of pu-rification. In the first, the solution can be com-pletely decopperized in a system of special lib-erator cells with insoluble anodes; arsenic andantimony are almost completely deposited andreturned to pyrometallurgical operations. Theelectrolyte is then concentrated by vacuumevap-oration to yield concentrated sulfuric acid andcrude nickel sulfate, fromwhich pure nickel sul-fate or nickel metal can be produced.

The second method of purification is by pro-ducing copper sulfate. For this purpose, the sul-furic acid is usually neutralized by addition ofcopper shot in the presence of air. The coppersulfate is obtained by crystallization, and themother liquor is cemented with iron scrap.

Anode Slimes [153], [154]. The content ofinsoluble substances is < 1 % of the anodeweight, and they collect on the bottom of cellsas anode slime. They contain precious metals(silver, gold, and platinum); selenides and tel-lurides of copper and silver; lead sulfate; stan-nic [tin (IV)] oxide hydrate; and complex com-pounds of arsenic, antimony, and bismuth (theundesired floating slimes). Themain componentis copper. In addition, gypsum and silica, alu-mina, or baryte from anode casting are present.

48 Copper

Table 18. Industrial range of copper anode and cathode compositions

Element Anodes (%) Cathodes (%)

Cu 99 – 99.8 99.99 + (excludingoxygen)

O 0.1 – 0.2 not determinedAg 0.01 – 0.5 0.0005 – 0.002S 0.0005 – 0.005 0.0004 – 0.001Ni 0.002 – 0.7 trace – 0.0003Pb trace – 0.5 trace – 0.0004Fe 0.001 – 0.006 0.0001 – 0.0002Sb trace – 0.1 trace – 0.0002As trace – 0.2 trace – 0.0002Se trace – 0.2 trace – 0.0001Te trace – 0.03 trace – 0.0001Bi trace – 0.03 trace – 0.0001Au trace – 0.004 trace

Table 19. Fractions of anode elements entering residues and elec-trolyte

Metal Percentage into anoderesidues

Percentage intoelectrolyte

Cu < 0.2 > 99.8Au 99 < 1Ag 98 2Se and Te 98 2Pb 98 2Sb 50 50As 30 70Co 5 95Ni 5 95Fe 0 ca. 100Zn 0 ca. 100Bi ∗

∗ Dissolves up to about 0.15 – 0.2 kgm3 in electrolyte, thenforms slimes.

The distribution of the elements (in %) variesover wide ranges:

copper 20 – 50nickel 0.5 – 2lead 5 – 10arsenic 0.5 – 5antimony 0.5 – 5bismuth 0.5 – 2selenium 5 – 20tellurium 1 – 4silver 25gold 4

Although the separation techniques differgreatly fromplant to plant, anode slimes are gen-erally treated as follows:

1) Oxidizing leaching of copper with dilute sul-furic acid

2) Recovery of selenium and tellurium by py-rometallurgical or hydrometallurgical meth-ods

3) Removal of detrimental elements and pro-duction of silver alloy (Dore bullion)

4) Separation of precious metals by electrolysis(silver and gold) and fractional precipitation(platinum metals)

The greatest part of the world selenium pro-duction comes from processing copper anodeslimes.

6.2.2. Practice of Electrorefining [155–157]

Electrorefining is carried out in large tankhouseswith capacities up to 500 000 t/a of pure cathodecopper. Today two techniques with principally

different starting cathode sheets are used. Theconventional system uses thin copper sheets,while the modern systems (ISA developed byCopper Refineries [167], Australia and Kidd de-veloped by Kidd Creek Divison, Falconbridge,Canada) utilize permanent stainless steel cath-ode sheets. The cathode copper is strippedmechanically from the steel sheets, which areused again, while copper starting sheets have tobe prepared for each cathode. The advantagesof the permanent cathode systems are high de-gree of automation, very low personnel costs,higher specific production rate, and higher cath-ode quality (lower content of impurities, e.g.,lead).

Cells and Operation. The cells or elec-trolytic tanks are constructed of reinforced con-crete lined with antimonial lead sheet or withflexible poly(vinyl chloride) or of polymer con-crete without a lining. The cells are rectangular,3.5 – 6m long, 0.8 – 1.2m wide and 1.2 – 1.5mdeep. Copper anodes and starting sheet cathodesare suspended alternately in the cells with pre-cise spacing by overhead crane. In most refiner-ies the anodes are prepared (pressed to makethem uniformly flat and weighed) before theyare inserted into the cells. Commercial cells con-tain between 40 and 60 electrodes of each type.The distance between the surfaces of the two ad-jacent electrodes should be small (about 2 cm)to avoid voltage loss, but not so small that thedeposited copper is contaminated with anodeslime. Electrolyte is circulated through each cellat ca. 0.02m3/min. At this flow rate the elec-

Copper 49

trolyte in one cell is exchanged every 5 – 6 h,which is important for removing impurities fromthe cell.

The anodes remain in the cell for 20 – 28 d.During this time the cathodes are usuallychanged two or three times, so that their weightis only 60 – 160 kg. The cathodes are removedby crane and the electrolyte on their surface iswashed off with hot water. In tankhouses whichuse stainless steel sheets, the plated copper isstripped mechanically (automatic stripping ma-chine) from both sides of the blank. In conven-tional tankhouses copper starter sheets are pre-pared by electrodeposition of copper on titaniumsheets (< 1mm, about 1 d) in special strippercells. Then the copper sheets are attached torolled copper support bars for hanging in thecells. This process is partly mechanized [164].

About 15% of the anode is not dissolved dur-ing the refining process. This anode scrap is re-cycled to the smelter (converter or shaft furnace).The anode slimeof the complete cycle is drained,filtered, washed, and sent to the precious metalplant. Part of the electrolyte is removed continu-ously from the circulation system for purifying.The sulfuric acid is sent back to the tankhouse.

Electrical System. The anodes and cathodesin a cell are connected in parallel. The cells areconnected in series to form sections of 20 to 40cells. Each section can be isolated electricallyfor inserting or removing anodes or cathodes.For example, 12 sections are connected with onesilicon controlled rectifier for current input. To-tal electrical current through the cell is between10 000 and 40 000A. Electrical densities of up to40 kA/m2 are used in ISA tankhouses today. Toobtain a high current efficiency of 92 – 98%, thebath and section voltages are monitored [254],[255], and the temperature of the cells is mea-sured by infrared techniques (from the crane). Inthis way short circuits, caused, for example, bythe growth of copper nodules can be quickly de-tected and corrected. The energy consumptionof electrorefining per tonne of plated copper isabout 260 – 280 kWh.

Electrocrystallization [163]. The structureof an electrodeposited cathode surface can beinfluenced by adsorption of molecules on thecrystallites. The addition of colloids or specialorganic compounds improves cathode quality by

yielding an evenly grained copper deposition.Important surface-active additives, or inhibitors,are bone glue, gelatin, avitoneA (sulfonatedaliphatic hydrocarbons), goulac, or bindarene(sulfonated wood fibers). Effective substanceswith definite composition are thiourea and itsderivatives and other sulfur – nitrogen com-pounds. These inhibitors, mostly in combina-tion, are added to the electrolyte in extremelysmall amounts, althoughmore is added at highercurrent density. Inhibitors increase the voltageand, therefore, the energy consumption. Severalrefineries automatically control their reagent ad-dition rates by measuring glue and thiourea, forexample, with the CollaMat [254–256] system.This stabilizes refinery operation and optimizesthe process (lower energy consumption).

Special Developments. Jumbo tanks insteadof conventional cells, first installed in 1972 atOnahama Tankhouse, Japan [165]. The jumbotanks have about 20 times the volume of con-ventional cells. The electrolyte flows parallel tothe electrode surface, in contrast to conventionalcells. A newer installation of jumbo tanks is theKidd Creek tankhouse in Canada.

To overcome anodic passivation caused byhigh current density and impure anodes a peri-odic current reversal system (PCR) [166] wasdeveloped. Optimum reversal conditions seemto be 20 units forward time and 1 unit reversetime [257]. This system has been adopted insome refineries, but due to the extra energy costsit is not widely used.

6.3. Melting and Casting

Copper cathodes must be remelted and cast toshapes because the structure of copper formedby electrocrystallization is not suitable forwork-ing to semifinished products. Cathodes areremelted in several types of special furnaces thatperform the tasks of melting, postrefining (ifnecessary), holding, and casting. Casting can becarried out by the older discontinuous methodsor by modern continuous casting.

50 Copper

6.3.1. Remelting of Cathodes

There are various kinds of furnaces that use ei-ther fossil fuels (coal, coke, fuel oil, natural gas,or reformed natural gas) or electric energy:

Small coke-fired crucible furnacesGas- or oil-fired rotary furnacesLarge hearth furnaces (reverbs)Electric-arc furnacesLow-frequency induction furnacesCathode shaft furnaces (e.g., ASARCO type)

Copper ready for pouring must be nearly freeof sulfur, at most 10−3% (10 ppm) S, because ahigher content affects detrimentally themechan-ical properties of the metal.

In practice, copper is treated in two ways:

1) After melting cathodes with sulfur-containing fuels, the copper melt must befire-refined like blister copper, by oxidationand poling. This is the case when hearthfurnaces are employed for casting.

2) Use of electric power or sulfur-free fuels al-lows the use of continuous units, such as in-duction or ASARCO furnaces.

The ASARCO shaft furnace, constructed byAmerican Smelting and Refining Co. [168],[169], has a cylindrical shaft consisting of a steeljacket with a brick lining. Cathodes are chargednear the top, and the combustion gases ascendin countercurrent flow from groups of burners;the liquidmetal is collected in a holding furnace.Apart from its effectiveness and high productiv-ity, a distinct advantage is the maintenance of aconstant, slightly reducing atmosphere by auto-matic control. The largest types (1.8m diameterand 8m high) can have a throughput up to 80 t/h.Worldwide ca. 200 units are in operation.

6.3.2. Discontinuous Casting

The discontinuous casting of various shapeson horizontal casting wheels with open ingotmolds, analogous to the casting of anodes, wasformerly themost important castingmethod. It isbeing replaced by continuous casting processes.

6.3.3. Continuous Casting

Since the end of World War II, several contin-uous casting methods have been developed. A

comprehensive synopsis, including patents, hasbeen published [171].

Continuous casting has several technologicaland economic advantages over the older castingprocesses, such as excellent surface quality, uni-form grain structure without blisters and shrink-age cavities, and energy savings.

Molten copper flows continuously into a rel-atively short water-cooled chill, usually linedwith graphite and open at both top and bottom.The solidifying metal itself forms the lower clo-sure. The shape is steadily withdrawn by clamp-ing rolls and cooled by spraying with water.

There are three main continuously castshapes: circular billets with a diameter up to500mm (for extrusion presses or tube rollingmills), square bars, and rectangular cakes withcross sections up to 1300× 200mm (for rollingto sheets and strips). Seamless tubes and otherhollow shapes are occasionally produced. Con-tinuous casting shapes are automatically sawnoff by flying saws when they reach the requiredlengths (up to 7.5m). In semicontinuous casting,the process must be interrupted, but the shapescan be up to ca. 12m long.

The kind of chill is generally independent ofthe furnace: the liquid metal stream free-fallsinto the chill (e.g., Junghans system), or the lessfrequent fixed connection of chill with the fur-nace (e.g., ASARCO casting system).

Other methods employ a joint moving in uni-son with the solidifying metal and a mold withtraveling parts forming a gap of the requiredcross section. (An example is the Hazelett pro-cess employed for continuous casting of anodes.Section 6.1.3.) Such casting methods are par-ticularly suitable for continuous production ofshapes with a small cross section. Section 6.3.4.

6.3.4. Continuous Rod Casting and Rolling

The continuous production of cast and rolledwire rod [172], [173] involves considerable en-ergy savings because the solidified hotmetal canbe rolled immediately. Several plants use such adirect process starting from molten electrolytictough-pitch copper. Worldwide >100 plants arein operation, using the following systems:

Copper 51

1) Properzi process [174]. The first continuouscopper rod caster was constructed in 1960following developments for other nonferrousmetals. It operates on the wheel – belt castingprinciple, i.e., casting into the gap betweenthe periphery of the casting wheel and theclosing steel belt. About ten plants are in op-eration with a capacity up to 30 t/h.

2) Southwire process [175]. Started in 1963 as afurther development of the wheel – belt cast-ing principle with capacity up to ca. 50 t/h,the Southwire process directly introducesthe continuous cast bar into the rolling mill(Fig. 33). After rolling, the rod, which is ox-idized on the surface, is continuously treatedby pickling with dilute sulfuric acid or alco-hols,water or steam rinsing, andwax coating.The saleable product (8 – 20mm diameter) isformed into “coils” of up to 10 t, which arepackaged. About 30 plants exist at present.

3) Secor process [177]. Only two factories(Australia and Spain) use this modifiedwheel – belt casting concept, dating from1975, with a capacity up to ca. 10 t/h.

4) Contirod or Krupp –Hazelett process [178].As a variant of the Hazelett twin-belt pro-cess similar to the Contilanod process (Sec-tion 6.1.3.), the continuous cast bar solidifiesbetween two belts and damblock chains andis directly moved to the rollingmill (Fig. 33).Metallurgie Hoboken-Overpelt, Belgium,developed this system in the 1960s; the ca-pacity of the largest units is ca. 50 t/h. Atpresent about 20 plants are operated.

5) General Electric dip forming process [179].A process based on the “candlestick” prin-ciple has been operated since about 1970. Acopper core rod is pulled upward through liq-uid copper so that its diameter increases; thethickened rod moves immediately to rolling.Oxygen-free copper can be produced by us-ing a reducing atmosphere. The capacity isca. 10 t/h. Nearly 20 plants exist at present.

6) Outokumpu up-cast process [180]. A newupward casting system developed in 1969draws copper upward through a vertical diecooler with a cooled graphite mouthpiecedipping into the melt. The caster, compris-ing 8 or 12 strands, yields oxygen-free cop-per at the rate of ca. 2 t/h per line. Because ofthe small cross section (8 – 25mmdiameter),

hot rolling is not required. Approximately 40plants are in operation.

6.4. Copper Powder

Copper and copper-alloy powders are requiredfor products prepared by powder metallurgicaltechniques, including friction materials, carbonbrushes, self-lubricating bronze bearings, spe-cial filters, and catalysts and other sintered com-ponents.

The principal methods for producing cop-per powders are electrolytic deposition at highcurrent densities and the atomization of moltenmetal, the latter more for copper-alloy powders.Copper powders are also formed by cementa-tion or by pressurized precipitation from aque-ous solutions, but such precipitates are of littlecommercial interest.

Atomizing is done by spraying a melt intoa pressurized air or water flow. Various grainshapes are formed, depending on the coolingrate and on additives that change the surfacetension. Additives that decrease surface tension,e.g., magnesium, form irregular powders; thosethat increase it, e.g., lead or phosphorus, yieldglobular particles. Spongy powders can be ob-tained by reduction of oxidized copper powderswith hydrogen.

Electrolytically deposited powder particleshave dendritic shape; a typical flow sheet isshown in Figure 34. For this purpose electrolysisis used as a shaping process rather than for refin-ing because high-purity copper cathodes are theanodes. The main parameters of powder elec-trolysis are, as Figure 35 shows, the following[181]:

Cathodic current densityElectrolyte temperatureConcentration of copper (II) ionsConcentration of chloride ions

Varying these factors markedly change particleshapes and apparent densities.

Anodic and cathodic current densities dif-fer. The former is normally 300 – 600A/m2, andthe latter is 2000 – 4000A/m2, which is 10 – 20times higher than the cathodic values in conven-tional electrolytic copper refining. This effect isobtained by using copper rod cathodes and plate-like anodes. The energy consumption in powder

52 Copper

Figure 33. Scheme of continuous rod casting and rolling [176]A) Southwire process; B) Contirod process.a) Melting furnace (ASARCO); b) Holding furnace; c) Wheel; d) Tundish; e) Steel band; f) Continuously cast copper bar;g) Preliminary rolling mill train; h) Finishing rolling mill train; i) Pickler; j) Coiler; k) Casting receptacle; l) Casting canal;m) Stationary edge dams; n) Middle rolling mill train

electrolysis averages nearly 2 kWh kg−1. Thepowders are generally posttreated for variouspurposes. Electrolytic copper powders are char-acterized by dendritic particle shape, high purity,low oxygen content, favorable resistance to ox-idation, and good green strength.

During the last 5 years the electrowinningtechnology to produce copper powder was de-veloped to commercial scale. By using pregnantleach solution from ore leaching operation andsubsequent solvent extraction and electrowin-ning in specially designed cells, it is possibleto produce dendritic shape copper powder withwell defined parameters like particle size [258].

6.5. Copper Grades andStandardization

The three main grades of refined copper are (1)tough-pitch copper, (2) deoxidized copper, and(3) oxygen-free copper.

Without regard to the method of refining,tough-pitch copper normally contains 0.02 –0.04wt%O (as Cu2O) and ca. 0.01wt% total ofother impurities. This grade is easilyworked andhas an electrical conductivity of 100 % IACS,but it is unsuitable for welding and brazing be-cause of the danger of hydrogen embrittlement.About 80 % of the world production of refinedcopper is tough pitch, mostly electrolytic tough-pitch copper (ETP).

Deoxidized copper, with no oxygen contentis produced by addition of nonmetallic or metal-lic reducing agents, usually copper phosphide.As a result of the absence of oxygen, hydrogenembrittlement cannot occur, and such copper iswell suited for welding and brazing. The con-tent of residual phosphorus, however, increasesthe electrical resistivity. Deoxidized copper withlow residual phosphorus (DLP, ca. 0.005% P) isrequired if the metal is to be used as a conduc-tor. Copperwith high residual phosphorus (DHP,

Copper 53

Figure 34. Flow sheet for electrolytic copper powder pro-duction [181]

Figure 35. Apparent density of electrolytic copper powderas a function of four electrolysis parameters [181]Other conditions: A) Cu 6 g/L, no Cl, 60 C; B) Cu 17 g/L,no Cl, 3600A/m2; C) no Cl, 50 C, 3600A/m2; D) Cu13 g/L, 50 C, 3600A/m2

ca. 0.04 % P) can be employed for nonelectricalpurposes.

Oxygen-free (OF) copper is a special qual-ity produced by keeping oxygen away from thecopper melt in a controlled atmosphere. Cop-per of this type with an oxygen content lessthan 0.001wt%, is suited for purposes requiringweldability and high electrical conductivity, es-pecially electronics. OFHC, oxygen-free high-conductivity copper, is internationally known.

Standardization [14], [182], [183]. Mostindustrial countries have established standardsfor copper; these national specifications includedetailed specifications for chemical composi-tion, physical properties, and geometrical di-

mensions, but differences exist. Table 20 indi-cates the rough equivalent between internationaland European specifications for the most impor-tant copper properties.

6.6. Quality Control and Analysis

Tests for quality control of copper are carried outon samples of both refinery shapes and semifin-ished products. There is need to standardize thetestingmethods, but currently only someof themare fixed in national specifications. Themost im-portant tests are the measurement of electricalconductivity, mechanical properties, and qual-ity of the metal surface.

54 Copper

Table 20. Comparison of international and selected national standards

InternationalISO R1337

European ENprEN1977prEN1978

MaterialsNumberprEN1412

Composition, % Electricconductivity, %IACS

Remarks

Cu min O2 max P min P max

Cu-CATH1 Cu-CATH1 CR001A 99.99 higher gradecathodes

Cu-CATH2 Cu-CATH2 CR002A 99.9 standard cathodesCu-ETP1 Cu-ETP1 CR/CW003A 99.9 0.04 101 electrolytically

refined, tough pitchcopper

Cu-OF1 Cu-OF1 CW007A 99.99 101 oxygen-free copperCopper not produced from Cu-CATH-1 Cathodes:Cu-ETP Cu-ETP CR/CW004A 99.99 0.04 100 electrolytically

refined tough pitchcopper

Cu-FRHC Cu-FRHC CR/CW005A 99.99 0.04 100 fire-refined toughpitch copper

Cu-OF Cu-OF CR/CW008A 99.95 100 oxygen-free copperCu-FRTP Cu-FRTP CR/CW006A 99.9 0.1 fire-refined copperPhosphor containing Copper:Cu-PHC Cu-PHC CR/CW020A 99.95 0.001 0.006 100Cu-HCP Cu-HCP CR/CW021A 99.95 0.002 0.007 98.3Cu-DLP Cu-DLP CR/CW023A 99.9 0.005 0.013 deoxidized copper

with low P contentCu-DHP Cu-DHP CR/CW024A 99.9 0.015 0.04 deoxidized copper

with high P content

The electrical conductivity is very sensi-tive to impurities and crystal lattice imperfec-tions. Mechanical tests measure hardness, ten-sile strength, elongation at failure, and torsionalfatigue strength. The spiral elongation test [184]is a complicated testmethod that assesses the pu-rity and the mechanical behavior of the sample.Defects on the surface and subsurface can occurin various forms, e.g., folds, splashes, cracks,inclusions, and voids. The voids are caused bygas porosity, shrinkage porosity, and shrinkagecavities.

Nondestructive testing procedures, such asradiography, ultrasonic examination, or theeddy-current technique can be used. Metallo-graphic methods (polishing and etching) giveinformation by microscopic examination.

Analytical methods are of interest for de-termining the impurity level of copper prod-ucts. Both wet chemical procedures and phys-icochemical procedures, such as atomic ab-sorption spectrometry, optical emission spec-troscopy, and X-ray fluorescence spectroscopy,are employed, the latter essentially for quickanalysis of solid and liquid co- and byproducts[185], [186]. The classical analytical methodsare gradually being superseded by the modernautomatic instrumental techniques. It is espe-

cially important to analyze the oxygen content,and one effective modernized method is avail-able, hot extraction, i.e., melting a copper sam-ple in a small graphite crucible and determiningthe CO formed by IR absorption spectroscopy.

7. Processing and Uses

The pure metal produced in refineries or remelt-ing plants is manufactured into semifabricatedproducts.

7.1. Working Processes

Usually copper is treated initially by noncutting,shaping processes to obtain semifinished prod-ucts or “semis”. These processes are subdividedinto hotworking, coldworking and, if necessary,process annealing.

Hot workingmeans plastic forming above therecrystallization temperature. Generally copperis preheated to 800 – 900 C, and the subsequenthot forming is finished at ca. 400 C. Cast barsfrom modern combined continuous casting/rod-rolling systems already have sufficient temper-ature, thus saving thermal energy. After cool-ing, the hot-worked copper is soft copper. Its

Copper 55

mechanical and electrical properties are scarcelychanged, but its density has increased to nearly8.9 g/cm3.

The next step is cold working, which involvesplastic forming below the recrystallization tem-perature. In practice, the operation is done atroom temperature. Unlike hot working, this pro-cedure entails an essential strain hardening ofthe metal by increasing the number of latticedefects; however, simultaneously formed latticevoids cause a considerable decrease of electricaland thermal conductivity. After cold working,the metal is hard copper.

Process annealing is a heat treatment that isnecessary if the hardened copper must be soft-ened again, either for continued working or forproducing (soft) copper with high electrical con-ductivity. Special furnaces are used for the pur-pose of steady heating and cooling of themetal –often in a nonoxidizing atmosphere. To achievethe intended microstructural change, the recrys-tallization temperature of 200 – 300 C must beexceeded; in practice, the metal is heated to400 – 500 C for accelerated recrystallization.Copper products with exactly defined propertiescan be obtained if all annealing conditions arecarefully controlled.

Table 21. Important fabricating processes for copper products

Refineryshapes

Hot-workingprocess

Cold-workingprocess

Semi-finishedproducts

Cakes hot rolling → cold rolling sheets/stripRod −→wire

drawingwires

Billets extruding → colddrawing

rods/sections

Billets extruding → colddrawing

tubes/pipes

or orhot rotarypiercing

→ cold pilgerrolling

The engineering techniques are versatile. Thefollowingworkingmethods are of special impor-tance:

hot working cold workinghot rolling cold rollingextrusion cold drawingdrop forging cupping

Foils only ca. 0.002mm thick are manufacturedby rolling, and wires to 0.004mm diameter by

drawing. Many products of varying size arefabricated by modern variants of the extrusionprocess [187]. The fabrication of tubes is alsoquite diverse [188]. Themost widely used work-ing processes are compiled in Table 21.

7.2. Other Fabricating Methods

In many cases, machining operations are re-quired, e.g., cutting, turning, planing, drilling,and sawing. However, these are more importantfor copper alloys than for pure copper becauseof copper’s tendency to gum. Noncontinuouscasting processes are likewise more suitable forcopper alloys because copper has a disadvanta-geous coolability. These include sandmold cast-ing, permanent mold casting, gravity die cast-ing, pressure die casting, and centrifugal cast-ing. Continuous or semicontinuous casting pro-cesses, however, are well-suited for pure copper.

Galvanoplasty is an electrolytic operation formanufacturing complicated objects that requirehigh precision and flawless surfaces such as hol-low bodies, disk matrices, and electrotypes. Aspecial galvanicmethod is copper plating,whichinvolves electrolytic deposition of a thin layer ofcopper on another metal either for surface pro-tection or as a base layer for electroplating withanother metal (→Electrochemical and Chemi-cal Deposition).

Powder-metallurgical techniques are usedprimarily for the mass production of smallpieces, especially intricate forms such as elec-trotechnical andmechanical structural parts. Themetal powders are first compacted by pressureand then sintered in a controlled atmosphere.The copper powder is often mixed with otherpowdered metals, including those that do notform common copper alloys (→Powder Met-allurgy and Sintered Materials).

There are other important fabricating meth-ods [189]. Joining is usually carried out aboveroom temperature by soldering, brazing, orwelding. Soldering may be used for all sorts ofcopper, owing to the low temperature. However,welding and brazing are feasible only with de-oxidized or oxygen-free copper.

When tough-pitch copper is heated in an at-mosphere containing hydrogen, the steamgener-ated (see page 7) collectswithin the grain bound-aries at high pressure and can destroy the grain

56 Copper

structure by forming cracks. This phenomenonis known as “hydrogen embrittlement.”

Mechanical joining and metal bonding arealso possible ways of joining copper with othermaterials.

Surface treatment of copper is a group ofoperations for surface protection or surface re-finement. These include mechanical, electri-cal, or electrochemical handling, e.g., polish-ing, matte finishing, pickling by dilute sulfu-ric or nitric acid, metal coating or electroplating(with nickel, nickel and chromium, tin, silver,gold, or platinum metals), lacquering or coat-ing with synthetic plastics (mainly for electricalinsulation), enameling of objects (applied art),and chemical or electrochemical coloring (dec-oration). Coloration is effected by chemicals,mostly specially formulated metal salt solutionswhich form thin layers of insoluble green, red,brown, or black compounds.

7.3. Uses

Copper is a useful material with a wide range ofapplications because of its combination of prop-erties. Because of its excellent electrical con-ductivity, it is the dominant conductor material.Copper is used primarily as round wire or rods,bare or insulated, for current generation, trans-mission, and conduction; various sorts of cablesare produced for special applications. Substan-tial quantities of copper are made into genera-tors, motors, transformers, and other electricalappliances. About 40 % of the world consump-tion of copper is for electrical purposes. As aresult of its high thermal conductivity, copper iswell-suited for vessels and pipes, especially forheating, cooling, and heat exchange.

While high-conductivity copper is requiredfor electrotechnical and electronic uses, specialcopper qualities are chosen for other uses. About30 % of world copper production is used for al-loying. Copper alloys are usually cold-worked;only ca. 10 % of them are cast.

Copper is frequently used in the chemical andfood industries because of its high resistance tocorrosion. There is substantial use of copper inmechanical engineering, by fabricators of pre-cision implements, in vehicle construction, andin ship building. There is increasing interest incopper building construction as a material for

installation, wall lining, and roofing. Hydraulicengineers use copper sheets for tightening ondams, sluices (floodgates), and bridges.

Other areas of application are in the fabrica-tion of household articles, art objects, coins andmedals, and inmilitary hardware as ammunition.There is a smaller demand for other purposes,such as electrodeposition; powder-metallurgicalcopper, special materials for brakes and self-lubricating bearings, small precision parts, fil-ters, graphite brushes; and alloying additives foraluminum, iron, and steel. Use in copper com-pounds, chiefly copper sulfate and copper ox-ides, consumes only 1 – 2%of the primaryworldproduction.

Table 22 lists the distribution of copper con-sumption among various industries.

Table 22. Industrial use of copper (including alloys) in the Westernworld in 1995, percentage by country [190]

Branch United Europe AsiaStates

Electrical and electronic industry 25 37.5 50Industrial machinery andequipment

11 9 9

Building construction 43 39.5 15Transportation 9 6.5 11Consumer and general products 12 7.5 15

Substitution and Miniaturization. Severalmaterials compete with copper and may sub-stitute for it, depending on the relative costs.Copper is partly replaced by aluminum in au-tomotive radiators and in transmission cables,high-voltage long-distance lines, and householdwiring. Copper wires and cables for telecom-munications are being displaced by microwavetechnology and fiber optics. Copper is beingreplaced by plastics for water pipes in both res-idential and commercial construction. In thearea of corrosion-resistant materials, in addi-tion to plastics there are also stainless steel andtitanium.

The movement toward making smaller andsmaller parts has been one of the most pervasiveand continuing pressures on the copper market.A dramatic drop in the use of copper has oc-curred in the widespread acceptance of printedcircuits. The use of wire has plummeted. Thenumber and size of the connectors have dropped.On the other hand, miniaturization has steadily

Copper 57

decreased the cost of the final products, thus in-creasing the number of units sold.

At the same time, however, this drive towardsminiaturization, whether in the thickness of anautomotive radiator or in the size of an electroniccomponent, is a challenge to the copper industryto produce purer copper and more useful alloysand to the copper fabricating industry to producethe new miniaturized products.

In several applications copper is resistingsubstitution by using new technologies. For ex-ample in telecommunications, copper continuesto be the preferred signal carrier for the last mile.The new xDSL (Digital Subscriber Line) tech-nology allows the existing copper infrastructureof ordinary telephone wires to also carry high-speed data. The installation of optical fiber incommunication trunk lines has led to a revolu-tion in the telecom industry. Copper applicationwas partly displaced, but overall this increasedthe demand for copper. Another development isthe use of copper circuitry in silicon chip tech-nology, which makes the microprocessors fasterand lowers energy consumption. Another exam-ple is the automobile radiator, which was for-merlymade of copper, whichwas then displacedby aluminum. New technology was developedfor producing smaller and lighter copper brassradiators with higher thermal conductivity thanaluminum radiators. A final example for innova-tion in copper is the development of supercon-ducting power cables made from high-tempera-ture superconductor wire. This technology willimprove energy efficiency, and now projects inChicago and Tokyo have been started.

8. Economic Aspects

There are numerous tabular compilations ofstatistics on copper resources, production, andconsumption [192–196]. Many books deal witheconomic relations and commercial problems,e.g., [197–200]. Compilations of companies inthe nonferrousmetal industry [201] and themin-ing industry [202] are published at irregular in-tervals.

Production and Consumption. The copperproduction ofmines, smelters, and refineries andthe consumption by country and region are givenin Table 23. Over the decades, the annual per

capita consumption of primary copper in theUnited States has grown [195]:

Late twenties 7.5 kgEarly thirties 2.5 kgWorld War II 9.5 kgPostwar period 7.5 kg1970 13.3 kg1979 14.6 kg

During the 1980s, there has been no increase inconsumption in America and Europe and only asmall increase worldwide. In the 1990s copperproduction increased due to industrial growth.

Figure 36 shows the development of worldper capita consumption for 1950 to 1997.

Cost of Copper Production and CopperPrice. The cost of copper production is char-acterized by high capital investment in miningprojects and in smelters and refineries. Miningprojects are financed by large consortiums andbanks. The capital investment for a green-fieldssmelter is in the region of 2500 – 3000 $/t ofdesign copper production. Smelter enlargementinvestment is approximately half that. There-fore, increased smelter production is preferablyachieved by enlargement.

In the last 15 years copper leaching projectshave been established. This is due to lower cap-ital investment than smelters (about half), andalso leaching operations built to increases thecopper yield of ores.

Operational costs are high due to energy con-sumption, which is the most important factor.For primary copper production the overall en-ergy consumption per tonne of copper is about45GJ, about half of which is consumed in min-ing and beneficiation and the rest in smeltingand electrorefining. For secondary copper, com-ing for example from copper scrap smelting andrefining, the overall energy consumption is only20GJ/t.

Due to several smelter enlargements, leach-ing operations, and also new energy-efficientmilling and smelting/refining processes, theoverall production cost of copper are falling.

The copper price is set primarily at the twometal exchanges: the London Metal Exchange(LME) and theNewYorkCommodity Exchange(COMEX). Like the copper production cost alsothe copper prices have been in an overall declin-ing trend since the World War II. The copper

58 Copper

Table 23. Production and consumption of copper in 1997 (in 103 t)

Mine Production Metal∗∗ Production Metal∗∗ Consumption

Argentina 30.2 15.6 51.7Armenia 1.9Australia 560.0 271.1 181.7Austria 77.0 31.0Belgium-Luxembourg 386.0 363.3Botswana 20.7Brazil 40.0 177.1 254.6Bulgaria 75.5 34.9 42.0Canada 658.0 560.3 224.6Chile 3 392.0 2 116.6 79.9China 495.5 1 179.4 1 258.0Colombia 1.8 4.0Congo 40.1 40.1Cyprus 3.9 3.9Czech Rep. 9.7Egypt 4.4 4.3France 35.1 558.0Georgia 4.0Germany 673.6 1 040.0Greece 96.8Hungary 13.3India 38.0 65.9 180.0Indonesia 548.4 54.2Iran 117.3 106.3 56.4Italy 85.7 520.6Japan 0.9 1 278.7 1 440.7Kazakhstan 316.2 301.1 13.7Korea (North) 10.0 30.0 15.0Korea (Rep.) 262.6 624.3Macedonia 13.0Malaysia 18.9 159.6Mexico 393.1 297.0 239.7Mongolia 130.0 3.0Marocco 15.4Myanmar 6.0Namibia 20.3Netherlands 40.4Oman 23.6Pakistan 11.0Papua New Guinea 111.5Peru 503.3 384.1 48.0Philippines 48.6 146.6 45.5Poland 414.7 440.6 233.0Portugal 106.5 1.4Romania 23.2 22.9 14.5Russian Fed. 526.0 610.0 165.0Saudi Arabia 1.0 145.0Scandinavia∗ 101.8 277.2 264.5Singapore 10.3Slovakia 23.0 28.0South Africa 185.6 130.2 81.8Spain 38.3 292.0 203.0Switzerland 7.5Taipei, China 587.8Thailand 151.9Turkey 65.0 113.7 194.4United Kingdom 60.4 408.5United States 1 940.0 2 452.4 2 790.0Uzbekistan 74.4 115.0 10.0Venezuela 18.0Yugoslavia Fed. 73.6 106.6 36.0Zambia 352.9 338.4 17.7Zimbabwe 6.8 17.9 15.0Others 1.8 16.0TOTAL 11 526 13 564 13 084

∗ Scandinavia includes Finland, Norway, and Sweden. ∗∗Refined copper.

Copper 59

Figure 36. Per capita consumption of refined copper, 1950 – 1997a) Relative per capita copper consumption; b) World populationBase index figures for 1950 are a copper consumption of 1.07 kg per capita and a world population of 2.5× 109

price also increases and decreases in the sameeconomic cycles as industrial growth and reces-sion. The development of copper prices since1960 is shown in Figure 37.

Product Information. Information onrecent developments in copper areavailable at http://www.copper.org andhttp://www.kupfer.org. Statistical data are avail-able at http://www.icsg.org.

9. Environmental Protection

The worldwide growth of industry and popula-tion has caused a series of environmental prob-lems, particularly the following: (1) emissioncontrol; (2)water protection; (3) solid-waste dis-posal.

Emission Control. There are two importanttasks in the treatment of off-gas from pyromet-allurgical processes in copper metallurgy: theremoval of sulfur dioxide and the containmentof flue dust.

Because most copper comes from sulfideores, sulfur is the main problem in copper ex-traction. In pyrometallurgical operations, it ap-pears as sulfur dioxide (Table 24) [204], [205].The mass ratio of sulfur to copper in sul-fidic concentrates is usually between 0.8 and1.6. Consequently, a large quantity of sulfurdioxide must be captured because of its harmful

effects on health, vegetation, and property. Inmost cases, sulfuric acid is produced from theSO2-containing off-gas (→Sulfuric Acid andSulfur Trioxide) [206].

Table 24. Comparison of SO2 concentrations in copper smeltingoff-gas [203]

Process SO2, vol %

Multiple-hearth roasting 5 – 8Fluidized-bed roasting 8 – 15Sinter roasting 1 – 2Blast furnace smelting 2 – 5Reverberatory furnace smelting 0.5 – 2.5Electric furnace smelting 0.5 – 5Outokumpu flash smelting 10 – 30INCO flash smelting 75 – 80KIVCET process 80 – 85Peirce – Smith converter 5 – 12Hoboken converter 7 – 17TBRC process 1 – 15Mitsubishi process 15 – 20Noranda process 10 – 20

Flue dust can be separated from off-gasto a high degree in modern gas-cleaning sys-tems such as electrostatic precipitators, bag-houses, cyclones, andwet scrubbers. Thismetal-containing dust is recycled.

Water Protection. Harmful wastewaterdoes not usually result from pyrometallurgi-cal copper production but water for direct orindirect cooling of furnaces, casting machines,and cast copper products is required on a largescale. This cooling water is moderately warmed,but does not acquire chemical impurities. Closed

60 Copper

Figure 37. Development of copper prices since 1960a) Cu price (current); b) Cu price (constant 1996)

circulation is used as much as possible in mod-ern plants.

Hydrometallurgical operations for the extrac-tion of copper from ores or concentrates presentthe risk of water pollution. These solutions, ofvarious compositions,must be posttreated if theycannot be recycled. Such posttreatment consistsof neutralization or precipitation of specific ions,chiefly anions bearing sulfur and the cations ofheavy metals. Lime is an excellent precipitant.Thus, the sulfate ion in sulfuric acid solutionsformed during hydrometallurgical extraction isprecipitated as gypsum [208].

Solid-Waste Disposal. The followingmeans are used for handling solid residues:

1) Recycling2) Exploitation as raw material for the prepara-

tion of useful products3) Dumping in deposits

One example of each method follows:

1) Flue dust from pyrometallurgical operations,e.g., from theOutokumpu flash smelting pro-cess, are added to the feed and recycled intosuitable furnaces (Section 5.5.1) and occa-sionally into blast furnaces after agglom-eration (Section 5.4.1). Zinc-containing fluedusts can be processed into zinc and zinccompounds.

2) Discarded slags with low copper contentfrom some copper smelting processes can besold after suitable treatment (Section 5.3.3).

3) Hydrometallurgical techniques yield variousprecipitates such as elemental sulfur or im-pure gypsum, which can easily be deposited.

10. Toxicology

Copper is a vital trace element for humans, mostanimals, and plants. The copper content of anadult human ranges from 50 to 120mg. The av-erage dietary intake of copper by adults rangesfrom 0.9 to 2.2mg/d. For higher organisms, thecompactmetal is completely harmless. HoweverProtista, especially bacteria, die in contact withmetallic surfaces of copper and many of its al-loys (oligodynamic effect) [209], [210].

Continued exposure to the metal dust orfumes can irritate mucous membranes. The fol-lowing exposure limits have been established:

Form Federal Republic ofGermany

United States

Metal MAK TLV/STELdust 1mg/m3 2mg/m3

Metal MAK TLV/TWAfumes 0.1mg/m3 0.2mg/m3

Copper 61

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