Energy Consumption in Copper Smelting: A New

JOM

DOI: 10.1007/s11837-015-1380-1 Ó 2015 The Minerals, Metals & Materials Society

Energy Consumption in Copper Smelting: A New Asian Horse in the Race P. COURSOL,1,5 P.J. MACKEY,2 J.P.T. KAPUSTA,3 and N. CARDONA VALENCIA4 1.—5N Plus Inc., Montreal, QC, Canada. 2.—P.J. Mackey Technology, Inc., Kirkland, QC, Canada. 3.—BBA Inc., Montreal, QC, Canada. 4.—Deltamet Consulting, Pointe-Claire, QC, Canada. 5.—e-mail: [email protected]

After a marked improvement in energy consumption in copper smelting during the past few decades, technology development has been slowing down in the Americas and in Europe. Innovation, however, is still required to further reduce energy consumption while complying with stringent environmental regulations. The bottom blowing smelting technology being developed in China shows success and promise. The general configuration of the bath smelting vessel, the design of high-pressure injectors, and the concentrate addition system are described and discussed in this article with respect to those used in other technologies. The bottom blowing technology is shown to be operating at a temperature in the range of 1160–1180°C, which is the lowest reported temperature range for a modern copper smelting process. In this article, it is suggested that top feeding of filter cake concentrate, which is also used in other technologies, has a positive effect in reducing the oxidation potential of the slag (p(O2)) while increasing the FeS solubility in slag. This reduction in p(O2) lowers the magnetite liquidus of the slag, while the increased solubility of FeS in slag helps toward reaching very low copper levels in flotation slag tailings. The application of high-pressure injectors allows for the use of high levels of oxygen enrichment with no requirements for punching. Using a standard modeling approach from the authors’ previous studies, this article discusses these aspects and compares the energy consumption of the bottom blowing technology with that of other leading flash and bath smelting technologies, namely: flash smelting, Noranda/Teniente Converter, TSL (Isasmelt [Glencore Technology Pty. Ltd., Brisbane, Queensland, Australia]/Outotec), and the Mitsubishi Process (Mitsubishi Materials Corporation, Tokyo, Japan).

INTRODUCTION One of the biggest business stories of 2014 was the huge drop in the price of oil. Thus, the West Texas Intermediate oil price dropped amid an oil surplus and lower demand by almost 50% from approximately $100/barrel in the middle of 2014 to around $50/barrel by early 2015. For a large energy consumer such as a copper smelter, this provided some relief to ever rising operating costs. However, analysts expect some rebound in the oil price later this year or into 2016. Hence, energy consumption in smelting, examined in this article for a number of

technology configurations, remains an important topic for copper smelters. The first published concept of bottom blowing smelting for nonferrous metals dates back to 1974 and the paper by Paul E. Queneau and Reinhardt Schuhmann1 titled ‘‘The Q-S Continuous Oxygen Converter.’’ The authors explained that the Q-S oxygen process invention was a response to the challenges of the time (first oil crisis resulting in high oil prices and pressure to fix sulfur dioxide gases) to increase process efficiency by a systematic use of oxygen with a substantial corresponding reduction in fossil fuel usage. Queneau and

Coursol, Mackey, Kapusta, and Valencia

Schuhmann adopted the following key concepts in their process design: Continuity—to limit capital and operating costs Autogenous—by using oxygen to lower energy and fossil fuel consumption Single off-gas of minimal volume—to lower costs of sulfur fixation Bottom blowing—to attain optimal solid–liquid– gas contact Countercurrent flow of matte and slag—to achieve bath oxidation and slag reduction in a single vessel An elongated, kiln-like vessel—to improve heat and mass transfer. In 1989, Queneau2 provided historical insights into the conceptual phase of the Q-S process development mentioning that a key driver that triggered the search for a new nonferrous smelting process was the invention of the Savard–Lee shrouded injector and its success in steel refining. In fact, Queneau and Schuhmann3 had filed their patent in 1973 and entered into a collaborative agreement with Savard, Lee, and Canadian Liquid Air that same year, forming QSOP Inc. (Queneau–Schuhmann Oxygen Process). With a lack of interest from the copper industry, QSOP found support from Werner Schwartz of Lurgi, leading to a QSOP-Lurgi agreement in 1974 for the development of the QSL (Queneau–Schuhmann–Lurgi) lead smelting reactor. After 15 years of efforts in the laboratory and in pilot and demonstration plants, including further developments of the Savard–Lee injectors, the QSL became the first bottom blown smelting vessel commercialized in nonferrous pyrometallurgy with lead smelting reactors installed in 1990 in Canada (Trail Smelter of Cominco Ltd.), Germany (Stolberg Smelter of Berzelius), and China (Baiyin Smelter of CNIEC), and in 1991 in Korea (Onsan Smelter of Korea Zinc). A more complete story of the QSL development from patent to commercial implementation was reported by Kapusta and Lee.4 The Chinese industry developed its own bottom blowing reactor in the 1990s. China Nonferrous Metal Industry’s Foreign Engineering and Construction Co. Ltd. and China ENFI Engineering (ENFI) first piloted their ShuiKouShan (SKS) lead smelting technology in 1999 at the ShuiKouShan lead smelter in Hunan Province. Commercial engineering and construction took place in 2001 and successful commissioning in 2002. The SKS copper process was first adopted in 2001 and commissioned commercially at the Sin Quyen Copper Complex in Vietnam in 2008 with a capacity of just 10,000 t/a of anode copper. The second implementation of the technology, and first in China, took place in 2008 at the Dongying Fangyuan Nonferrous Metals Co., Ltd. (Dongying) with an original copper concentrate smelting capacity of 32 dry t/h or 55,000 t/a of anode copper.5,6

The number of SKS lead reactors in China quickly grew after 2002; Stephens7 reported that the SKS lead furnace had been described ‘‘as the smelting section of a QSL reactor.’’ As for the SKS copper furnace, Kaixi Jiang et al.5 described it as similar in design to the Noranda Reactor, contrasting with the fact that ‘‘the air in the copper matte layer is blown into the furnace via the oxygen guns set in the furnace bottom.’’ These oxygen ‘‘guns’’ are essentially Savard–Lee-type shrouded injectors with compressed air as shrouding gas. The process at Dongying is now known as the bottom blowing smelting (BBS) process. Several other copper smelters in China have since adopted the SKS/BBS technology. It has also been reported that the technology is being evaluated as an alternative for some copper smelters in Chile. The goal of this article is to discuss the SKS/BBS technology features and to evaluate the energy requirements for this technology compared with other modern smelting technology. To do so, the approach used by Kellogg and Henderson8 and Coursol et al.9,10 is used. In this approach, all technologies are compared on the same basis, with the same concentrate, flux and coal composition, this allows evaluating both the electrical and thermal energy required to operate a smelter from concentrate to anode for a given technology. BOTTOM BLOWING SMELTING TECHNOLOGY (SKS/BBS) General Description The SKS/BBS reactor, as shown in Fig. 1, is a cylindrical vessel with gravity top feeding of wet concentrate through several ports. The injectors are all located on one side of the furnace, whereas the off-gas mouth, matte, and slag tap holes are on the opposite side. This arrangement creates a two-zone bath: an agitated oxidation zone below the feed ports and a quiescent settling zone above the matte tap hole. Two auxiliary burners are located on each end wall and are used during start-up or stand-by. The mouth is small compared with existing bath smelting reactors because operating at high oxygen

Fig. 1. Schematics of SKS/BBS copper process from Yao and Jiang.11

Energy Consumption in Copper Smelting: A New Asian Horse in the Race

enrichment produces lower volumes of process offgases. Cui et al.12 provided a complete description of the bottom blown oxygen smelting process at Dongying. The cylindrical furnace is 4.4 m in diameter by 16.5 m in length. The oxygen ‘‘lances’’ are positioned in two rows at the bottom of the furnace with five lances in the lower row at 7° to the vertical and the four lances in the upper row at 22°. This provides a 15° angle between the two rows. A photograph of the Dongying furnace is given in Fig. 2. The oxygen lances or guns used to inject the blast are Savard–Lee-type shrouded injectors with similar design and configuration as the original QSL injectors developed by Lurgi and Air Liquide in the 1970s.4 The tips of the injectors are built with a gear-like design as shown on Fig. 3a (drawing) and Fig. 3b and c (photographs) for the reactors built by ENFI. The principle of the lances is to inject highoxygen-enriched air or pure oxygen in the central bore and first or two annuli of grooves and a cooling medium in the outer ring of grooves (e.g., air or nitrogen). All streams are injected at sonic velocity. The lance design at Dongying has been slightly modified to reduce the pressure requirement of the oxygen streams to achieve sonic velocity.13 The oxygen enrichment achieved to date with the SKS/BBS reactors is in the range 50–75%, which is significantly higher than for the current Noranda

reactor and Teniente converter operations. The original design capacity of the Dongying furnace was 55,000 t/a of anode copper at 55% oxygen, and its current capacity has reached 100,000 t/a as the oxygen enrichment reached 75%. The inherent low level of nitrogen in the blast is frequently perceived as a weakness of high-oxygen injection due to the diminished mixing of the bath. A comparison with the operating oxygen top-blown– nitrogen-bottom stirred vessel at Vale’s Copper Cliff smelter in Sudbury is opportune at this point. Marcuson, Diaz, and Davies reported the use of five porous plugs, each with nitrogen flow rates of 14 Nm3/h.16 The total flow rate of nitrogen for stirring was 70 Nm3/h, which was evidently sufficient to provide stirring in a 135-t semiblister bath (corresponding to a specific blowing rate of 0.52 Nm3/h/t of the melt). In this later case, the top blown oxygen flow rate was approximately 5330 Nm3/h. With an SKS/BBS furnace operating at 75% oxygen, the total blast rate is approximately 19,160 Nm3/h, corresponding to the Dongying vessel to a specific blowing rate of 70 Nm3/h/t of melt. These conditions provide nitrogen in the order of 4,790 Nm3/h for stirring, a sizeable amount compared with the porous plugs in the Copper Cliff case. By comparison, the Noranda Process reactor in Canada has a specific blowing rate of about 130 Nm3/h/t of melt. The lance life was reported in 2013 by Xiaohong Hao et al. to be 30–60 days14 and in 2014 by Johnny Zhang et al. to be up to 6 months.17 This discrepancy highlights the importance of a proper lance design to ensure sonic velocity is achieved, and a controlled bath chemistry and temperature is needed to limit lance and refractory wear. Dongying has achieved the most advanced developments in the operation of its SKS/BBS furnace. This operation has therefore been selected in this article as the reference for energy comparison of the SKS/BBS furnace to the Noranda and Teniente reactors. Process Chemistry

Fig. 2. SKS/BBS furnace at Dongying (Source: Yao and Jiang.11)

Although most modern smelting vessels accomplish a similar task of producing a high-grade

Fig. 3. (a) Drawing of lance tip from Hao et al.,14 and (b, c) Photographs from July 2013 ENFI presentation.15


matte, the chemistry can vary significantly from one technology to another. In the BBS furnace case, several factors seem to affect positively slag chemistry, leading to low copper levels in the furnace slag and favorable oxygen consumption figures. A few of these aspects are discussed next. Impact of Matte Grade The smelting furnace matte grade is important for the process chemistry. A matte grade higher than 77% leads to higher soluble copper in slag in the oxidic form, whereas operating at a matte grade lower that 70% leads to moderate copper levels in slag, predominantly in the form of soluble sulfidic copper.18 In the BBS process, the operating matte grade is in a range where the lowest soluble losses are observed, which is generally between 70 and 75% Cu.18 A reasonably low total copper in slag is observed in the BBS furnace, indicating less than 5 wt.% matte entrained in slag.6 Given that the approach taken for slag cleaning is by milling and flotation, a high copper recovery is expected, with a copper content in tailings reported at 0.26%.6 FeS Solubility in Slag and Impact on Copper Recovery The bulk sulfur content in slag of the BBS reactor was reported to be 1.7 wt.%,6 whereas the bulk copper content of the slag was reported to be 2.9 wt.%.6 These numbers appear low from the authors’ viewpoint and are expected to vary as a function of slag quality and matte grade. If one assumes all copper is present as matte entrained in the slag, then the %Cu/%S ratio should be maintained at 3.4. If a significant level of copper oxide is present in the slag, then the ratio should be higher. In the BBS furnace, the %Cu/%S ratio is 1.71, indicating that another sulfide species is soluble or entrained in the slag. Previous publications have discussed FeS solubility in copper smelting slag and its significance.18,19 Figure 4 shows the sulfur solubility

0.7

Fe/SiO2=1.6 Fe/SiO2=1.4 Fe/SiO2=1.8

%S in Slag as FeS

0.6 0.5 0.4 0.3 0.2 0.1 0.0 0.5

2.0

3.5 5.0 6.5 %Fe in matte

8.0

9.5

Fig. 4. Calculated sulfur solubility (as FeS) in the Teniente converter slag. Effect of Fe/SiO2 ratio and %Fe in matte at T = 1250°C. (p(SO2) = 0.25 atm, [Al2O3]slag = 4.0 wt%, [ZnO]slag = 2.1 wt.%, [MgO]slag = 0.8 wt.%, [CaO]slag = 0.8 wt.%).19

trend as function of the matte grade for different Fe/ SiO2 ratios in slag from a bath reactor such as the Teniente Converter. The thermodynamic calculations assume all sulfur is soluble as FeS, but the copper sulfidic dissolution is yet to be modeled appropriately. At 70% matte grade and a Fe/SiO2 ratio of 1.8, the expected %S soluble in slag is 0.7 wt.%, which is lower than the level of 1.1 wt.% reported by Zhao et al.6 for the BBS slag, which must account for both Cu and Fe sulfidic dissolution. The sulfur solubility in slag can have a positive impact by forming recoverable matte droplets during slag cooling and solidification, if the slag is solidified at a sufficiently low cooling rate.20 Figure 5a shows the microstructure of a rapidly cooled slag sample, in which sulfide exsolution led to a phenomenon called ‘‘copper fog.’’ In the copper fog formation, the small sulfide particles are trapped between olivine and terminal glass (final slag to solidify in a glassy state during the solidification process) and are quite hard to recover by milling and flotation. On the contrary, when the slag is solidified under a controlled slow cooling rate, as practiced by several smelters worldwide, this results in a much coarser microstructure and the growth of droplets (e.g., due to coalescence phenomena). Hence, higher Cu recoveries can be obtained. Figure 5b shows a coarse microstructure with matte droplets formed during cooling slag and a large entrained matte droplet with sulfide exsolution texture. Feeding on Slag Instead of Submerged Injection In the BBS furnace case, the fresh feed is added as a wet concentrate onto the slag surface. Under these conditions, it is conceivable to obtain a lower p(O2) in the slag than in the matte because of the addition of fresh concentrate from the top, as long as mixing is adequate, yet not overly intense. With more intense mixing, the p(O2) of both phases would be nearly the same (closer to equilibrium). Melting of the concentrate within the slag can contribute to maintaining the dissolved copper oxide content to quite low levels and, furthermore, to controlling the magnetite level of the slag. In the BBS furnace case, the reduced bath agitation due to the higher oxygen enrichment and lower blowing rates (or, a lower total off-gas volume expressed as m3/m3 of melt) can provide favorable conditions in the slag to minimize the slag liquidus (lower p(O2) and a higher level of soluble FeS). Although the Teniente Converter with submerged injection generally operates near 1225– 1250°C, the Noranda Reactor and the Isasmelt furnace (Glencore Technology Pty. Ltd., Brisbane, Queensland, Australia) using wet feed addition from the top operate at 1200°C and 1180°C, respectively, which can lead to a significant improvement in the refractory protection and heat balance of these vessels. The BBS technology, operating with top feeding and high oxygen enrichment levels, seems to push


Fig. 5. (a) Fine microstructure showing ‘‘copper fog’’ (bright dots). (b) Coarser microstructure obtained by slow cooling of the slag, leading to high copper recovery by flotation (Cu2S and Cu-Fe sulfide [bright phases], fayalite [gray columnar crystals], terminal glass [dark gray phase between the fayalite blades], and magnetite crystals [intermediate grey crystals shown in (b)]).

the concept observed in other top feeding technologies toward a reduced mixing intensity, leading to even lower operating temperatures, reported to be in the 1160–1180°C range.6–17 Operating at High Oxygen Enrichment Higher oxygen enrichment has the benefit of releasing more heat during the smelting process and of allowing more reverts or low-grade materials to be recycled in the smelting vessel. It also has the advantage of providing a lower mixing level in the smelting vessel and reducing the off-gas volume. One should also note that these conditions reduce the risk of foaming during periods when the slag contains excessive levels of solid magnetite. One drawback of using high oxygen enrichment is the increase in OPEX and CAPEX for oxygen production. Oxygen Usage in Copper Smelting From practical experience with modern smelting technologies, the oxygen usage factor (or oxygen efficiency) in most technologies is in the 95–99% range. Although much lower oxygen usage per tonne of concentrate can apparently be observed in some cases, clearly an oxygen efficiency over 100% is not possible. Differences can be explained by the factors discussed next. It is noted that a lower specific oxygen consumption was observed in the BBS furnace case.17 When concentrate is fed to a smelting unit from the top (Noranda Reactor, El Teniente Converter, BBS furnace, etc.), the freeboard air or gas can react with the concentrate if sufficient oxygen is available. This phenomenon is illustrated by reactions 1 and 2. The chalcopyrite, bornite, and chalcocite present in the concentrate can be oxidized following similar mechanisms. Although this is not the major oxidation mechanism, it can make a significant difference in specific oxygen consumption for a given feed. The concentrate mineralogy obviously has a significant impact on the oxygen requirements. Mainly, the input FeS2 and FeS equivalent are very

important, as most of the remaining iron sulfide needs to be oxidized during converting. This is illustrated by reaction 3. The mineralogy of reverts added to the smelting unit can also lower tonnage oxygen usage and SO2 generation. For example, adding metallic copper in the form of reverts, copper scrap, or spent anodes can reduce both the SO2 produced during smelting and tonnage oxygen usage. This can be referred to as ‘‘sulfur sequestration’’ and obviously has an impact on the FeS oxidation mechanism, changing from reaction 3 to 4. Similarly, Cu2O contained in reverts from the converter aisle or anode furnaces represent a source of oxygen and therefore can also significantly reduce tonnage oxygen consumption. This is illustrated by reaction 5. Finally, in a deficient heat balance situation, coke addition is used for closing the heat balance. Most of this coke entering the smelting unit exits as CO2 gas with trace levels of CO; hence, the oxygen required for this reaction contributes to the total specific O2 consumption. ðFeS2 Þfeed þO2ðfreeboardÞ ¼ ðFeSÞmatte þSO2

(1)

ðFeSÞfeed þ3=2O2ðfreeboardÞ ¼ ðFeOÞslag þSO2

(2)

ðFeSÞmatte þ3=2 O2 ¼ ðFeOÞslag þSO2

(3)

ðFeSÞmatte þ2ðCuÞmatte þO2 ¼ ðFeOÞslag þðCu2 SÞmatte (4)

ðFeSÞmatte þCu2 O ¼ ðFeOÞslag þðCu2 SÞmatte

(5)

The preceding discussion is helpful in understanding the lower published and observed oxygen


use in the BBS furnace compared with the Noranda–Teniente reactor as indicated by Zhang et al.17 METHODOLOGY-ENERGY CALCULATIONS The methodology adopted to evaluate smelting energy in this study was based in part on the approach used by Kellogg and Henderson,8 which was reviewed for modern smelting technologies by Coursol et al.9,10 In these last two studies, the authors reviewed the energy consumption (electrical and thermal) and compared modern technology performances in specific configurations, process copper concentrate, and finished anodes. In total, 12 different flowsheet configurations were compared. A Metsim model (METSIM; Proware, Tucson, AZ) previously developed by Tripathi et al.21 was used as a basis for designing several other models and for comparing technologies on the same basis. Heat and mass balance and energy consumption data were then computed for each process for subsequent analysis and comparison. Process ‘‘boundary limits’’ were as follows: Inputs: wet concentrate, flux, and other consumables delivered to the smelter day bins; and Outputs: copper anodes, sulfuric acid, acid plant tail gas, cleaned fugitive gas released to atmosphere, and cleaned slag. Waste heat recovery from process gas streams was included as part of the study. In this work, the same approach was used to compare the BBS technology with other modern technologies. Energy consumption in auxiliary unit operations and energy equivalent for process supplies were computed from a set of unit energy consumption factors. These are presented in Table I based on data from Refs. 9 and 10.

Appropriate data were also taken from Kellogg and Henderson,8 and updated information was used in other cases. Examples of auxiliary unit operations include producing tonnage oxygen, compressing Peirce Smith Converter (PSC) injection air, delivering low-pressure air to burners, moving process off-gas, drying concentrate and other materials, and transporting and injecting fine solids suspended in a stream of gas (dense phase transportation of solid particulates). The standard conditions used for the calculation of energy requirements for each of the chosen processes are presented in Table II and are identical to the ones taken in previous studies.9,10 These processes include the assay of a standard copper concentrate and the data relative to fluxes and fuels.

Table II. Standard conditions used in the Metsim model Item and data Concentrate analysis (dry basis): 30.5% Cu, 28.5% Fe, 31.5% S, 5% SiO2, 2% Al2O3 Concentrate moisture content: 10% H2O Flux analysis (dry basis): 88% SiO2, 2% CaO, 6% Al2O3, 2% MgO, and 2% Fe3O4 Flux moisture: 3% H2O Natural gas: 37.3 MJ/Nm3 Coal: 28.4 MJ/kg Ambient conditions: 0°C, 760 mmHg

Table I. Unit energy parameters used in this study Item Steam dryer Conversion of steam to electricity at smelter Tonnage oxygen production (300 kPa) Tonnage oxygen production (600 kPa) Compress tuyere air (600 kPa) Compress tuyere air (110 kPa) Compress lance air (60 kPa) Process off-gases handling Fan-secondary gases Furnace cooling water Matte granulation and handling Slag granulation and handling Matte comminution and handling Lighting and miscellaneous power (allowance) Acid plant operation (double contact) Energy-Flux Energy-limestone calcination (CaO flux) Energy-wear steel in slag milling Energy-pig iron a

Unit energy

References

2 t steam/t water evaporated 6.25 kg of steam/kWha 285 kWh/t of oxygen 321 kWh/t of oxygen 0.126 kWh/Nm3 0.05 kWh/Nm3 0.03 kWh/Nm3 0.0085 kWh/Nm3 0.002 kWh/Nm3 3 kWh/t Cu 9 kWh/t of matte 3 kWh/t of slag 10 kWh/t of matte 30 kWh/t of Cu [(646.8/%SO2) + 63.7] kWh/t of acid (90 MJ + 3 kWh)/t of flux 7000 MJ/t of CaO 20.7 MJ/kg of steel 15.5 MJ/kg of pig iron

9 9 22 Authors Authors 8 9 9 24 9 9 9 9 8 8 8 9 23 9

Based on a rate of 5 kg steam/kWh and an operational efficiency of 80% to account for potential losses on start-up/standby, etc.


Brief Description of Processes Assumptions (Noranda Reactor Versus BBS Technology) Each process route also included the following ‘‘standard’’ unit operations: (I) complete secondary gas collection and cleaning, (II) anode refining and casting, and (III) process gas treatment in a doublecontact acid plant with acid delivery to storage tanks. As noted above, heat recovery from process off-gases was also used. Concentrate and other solid process streams requiring drying were treated in steam dryers using waste heat steam. Surplus steam was assumed to generate electricity. The flowsheet configuration used for the BBS process was identical to the configuration used for the Noranda Process by Coursol et al.10 Figure 6 shows the flowsheet configuration used in both the Noranda Reactor and the BBS furnace cases for the mass balance calculations. In both simulations, Noranda and BBS, a feed rate of 126.8 t/h of a standard filter cake concentrate was assumed (10% moisture, 114 t/h dry basis), were fed to the smelting vessel. The total heat losses for both smelting units were fixed at 7 MW. The two flowsheet had the exact same configurations as shown in Fig. 6. A few differences in process data between the two cases are listed next. In the Noranda Reactor case, with normal tuyere technology (low pressure and nonshrouded), the oxygen enrichment is limited to 45 vol.% to minimize risks on tuyere line integrity. The maximal coke addition to the vessel is limited to approximately 2 t/h from practical experience. In this simulation, the Fe/SiO2 ratio in the slag was set to 1.42. The final copper losses in slag tailings were assumed to be 0.39 wt.% at a slag concentrate grade

of 37 wt.% Cu.25 With an oxygen enrichment of 43%, the heat balance was closed with a slag temperature of 1244°C. These conditions, given as a reference, are the same conditions as presented by Coursol et al.10 In the BBS furnace case, with shrouded tuyeres, it is possible to achieve oxygen enrichment as high as 75 vol.%. By using this technology allowing for higher oxygen enrichment, the coke addition in the smelting unit was reduced to zero, and the slag concentrate grade was reduced to 15 wt.% Cu allowing a lower %Cu in slag tailings (0.26 wt.%) to be assumed. The lower grade of slag concentrate provides a greater tonnage of cold material to be added, thus being available to absorb excess smelting heat as a consequence of the relatively high level of oxygen enrichment used. Thus, with the two changes mentioned above, an oxygen enrichment level of 63% is allowed to close the heat balance and obtain a slag temperature of 1180°C. RESULTS Tables III and IV show the results of energy and fossil fuel consumption for the Noranda Reactor and the BBS furnace cases, respectively. Separate columns for electric energy, expressed in kWh/t of anode copper, and fuel, expressed in MJ/t of anode copper, are provided in these tables. The fuel equivalents of electrical energy were calculated using a power plant efficiency of 38%.9 In the tables, the numbers for items such as fuel, oxygen, compressed air, secondary gases, and fugitive gases correspond to respectively overall smelter consumption or production.

Fig. 6. Model flowsheet configuration used for the Noranda flowsheet and the BBS flowsheet.


Table III. Energy requirements for the Noranda reactor case with Peirce–Smith converters and slag flotation (slightly modified from Ref. 10 for consistency with following BBS furnace case) Electrical Item Fuel Tonnage oxygen High pressure air Process gas handling Secondary and fugitive gas handling Supplies, steel, etc. Milling slag Acid production Lighting and miscellaneous Steam credit

kWh/t

Eq MJ/t

0 256 135 61 76 2 160 401 30 1120 176 944

0 2427 1276 576 720 17 1515 3800 284 10,614 1669 8946

Fuel MJ/t

Total MJ/t

4088 0 0 0 0 39 0 0 0 4127

4088 2427 1276 576 720 56 1515 3800 284 14,741 1669 13,072

Table IV. Energy requirements for the BBS furnace case with Peirce–Smith converters and slag flotation Electrical Item Fuel Tonnage oxygen High-pressure air Process gas handling Secondary and fugitive gas handling Supplies, steel, etc. Milling slag Acid production Lighting and miscellaneous Steam credit

kWh/t

Eq MJ/t

0 298 133 48 76 1 160 360 30 1107 129 978

0 2824 1255 459 720 12 1518 3413 284 10,486 1223 9263

Fuel MJ/t

Total MJ/t

2331 0 0 0 0 39 0 0 0 2370

2331 2824 1255 459 720 51 1518 3413 284 12,857 1223 11,634

The calculations performed in this study and selected data from previous studies9,10 are shown in Table V.

Table V. Comparison between energy consumption for copper production (concentrate to anode) for the reverberatory furnace and for selected modern copper smelting technologies Processing route KH-hot calcine reverberatory8 Flash smelting–flash converting–slag flotation9 Isasmelt–Peirce–Smith converting-rotary slag cleaning9 Mitsubishi Process (Mitsubishi Materials Corporation, Tokyo, Japan)9 Noranda–Teniente with dry feed + slag flotation9 Noranda reactor (filter cake) + PSCs + slag flotation Bottom blowing smelting (filter cake) + PSCs + slag flotation a

All energies are expressed in MJ/t of anode copper.

Electric energya

Fossil fuela

Totala

2173 9266 6903 8508 10,088 8946 9263

15,935 1518 4175 2498 2657 4127 2370

18,108 10,784 11,078 11,006 12,746 13,072 11,634


CONCLUSION From an energy perspective, our calculations indicate the BBS/SKS technology to be superior to the Noranda/Teniente smelting vessels and to be nearly equivalent to other highly efficient technologies: flash smelting, Isasmelt, and the Mitsubishi Process. The calculated specific energy data presented in Table V are not only for smelting but also for treating concentrate through to anode copper; hence, the results of all technologies can be improved by optimizing peripheral equipment. The authors believe that more work needs to be done regarding better converting technologies and anode furnace design to further reduce energy usage. Oxygen enrichment, a dominant aspect in improving energy efficiency, has been incorporated in most modern technologies, and the industry now has to search elsewhere while reaching even higher enrichment levels. Converting, fire refining, waste heat recovery, and oxygen/acid production are all key areas to allow further improvements in energy efficiency and energy usage. The Noranda Reactor and Teniente Converter have been work horses in Canada and Chile during the last 40 years. Much attention has been placed in increasing productivity but with no major change in vessel configuration or injector technology. The current practical limitation in oxygen enrichment with low-pressure, nonshrouded tuyeres is considered near to 45% O2. It is believed that high-pressure injectors and/or shrouded tuyeres added to the Noranda–Teniente Converter can provide significant benefits in terms of energy efficiency, with attendant improvements in environmental performance. Using this approach, such vessels located in Chile could improve operational and environmental performance and better energy efficiency with likely a low capital investment. ACKNOWLEDGEMENT The authors would like to acknowledge the contribution of Dr. Carlos Diaz in previous publications on energy efficiency, which paved the way for the present study. REFERENCES 1. P.E. Queneau and R. Schuhmann, JOM 26, 2 (1974). 2. P.E. Queneau, JOM 41, 30 (1989). 3. P.E. Queneau and R. Schuhmann, U.S. patent 3 941 587 (2 March 1976). 4. J.P. Kapusta and R.G.H. Lee, Proceedings of Copper 2013, International Copper Conference, Vol. III (Book 2)—The Nickolas Themelis Symposium on Pyrometallurgy and Process Engineering, ed. R. Bassa, R. Parra, A. Luraschi, and S. Demetrio (Santiago: The Chilean Institute of Mining Engineers, 2014), pp. 523–558. 5. K. Jiang, L. Li, Y. Feng, H. Wang, and B. Wei, Proceedings of the T.T. Chen Honorary Symposium on Hydrometallurgy, Electrometallurgy and Materials Characterization, ed. S. Wang, J.E. Dutrizac, M.L. Free, J.Y. Hwang, and D. Kim (Warrendale, PA: TMS-AIME, 2012), pp. 167–176.

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