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By John V. Beall, William F. Haddon

A long the mighty Andean Cordillera, there is splendor beyond imagination-in the natural beauty of the mountains and in daring engineering and lavish investment in the mines. This is the story of the great copper mines of Chile and Peru. Each mine has a special quality which has tested the physical and spiritual resources of the founder. In the early days of the older mines, the mining engineer of song and story was more the rule than the exception. He was physically tough, ingenious and a leader of men. In the current political and social environment of Chile and Peru, the foreign contract mining engineer is still needed; but he must be a different type of man. To John J. Selters at Cerro's Rio Blanco, the principal ingredient is still "toughness," but it is "toughness of mind" that is needed. "He probably should be a teacher." The copper mines are not new. They were, almost without exception, worked by Indians (La Exotica might prove the rule), and later by the intrepid Spaniards. Modern mining began after the turn of the century. In Chile, the first of the big ones was the Braden Copper Co. founded on the El Teniente deposit. This development was followed by Chile Exploration Co. at Chuquicamata and Andes Copper at Potrerillos. These properties have been developed over the years by vast expenditures of money and a monumental amount of labor. None was free of technical problems-metallurgy at Chuqui, marginal grade of the old mine at Andes, avalanches and the great fire at El Teniente. The evolution which has brought the established mines of Chile, and the new ones coming on stream, to their present state of development is in reality the work of wonders of input-money, engineering and responsible management in an atmosphere of political stability. Without any one of these, the result would have been failure.

Jan 11, 1969

By John A. Dresser

Seventy years ago a period of prospecting and mining activity began in the Eastern Townships of Quebec that seems to have been quite equal to that of recent years in the Rouyn field. In the following six or seven years a large amount of work was done over a widespread area. About fifty deposits gave some shipment of ore, from a few tons upwards, and as many as five hundred occurrences of copper were made known in the district by official records. Comparing the scale of commercial undertakings of that time and of the present day, these developments must have occupied quite as large a share of the capital and man-power of the Province as northern Quebec employs today. Working conditions at that time were not very favourable. The country was then only newly opened for settlement, and colonists were few and widely separated. Roads of the class built today were unknown. A single line of railway, the Grand Trunk, crossed the district from east to west. Colonization companies controlled large blocks of land on which they owned mining rights and usually reserved these rights in whole, or in part, when selling the lands. Then, as now, the precious metals belonged to the Crown. Yet the work was not without its successes. One of the first deposits discovered, the Acton, was, for a few years, one of the most profitable producing copper mines of its time. Another, the Eustis mine, has been in almost constant operation ever since, and is now vigorously pushing further development. At least two others, now quiescent, proved profitable operations, and several made considerable production. The manufacture of sulphuric acid in Canada was begun by the founder of the Nichols Company at Capelton, on a deposit that is now being re-developed, after a long and creditable career.

Jan 1, 1928

By J. G. LECKIE

During the first ten months of 1943 copper was produced at a higher rate than in 1947. However, on Oct. 24 one of the large mines was shut down due to a strike. As of Dec. 31 the strike was still in effect and had already resulted in a loss of production of 50,000 tons of copper. The demand for copper was considerably greater throughout the year than the supply. The price of electrolytic copper delivered at the Atlantic Seaboard refineries advanced from 21.20 to 23.20 cents per lb during the year. This sustained demand for copper, together with the increase in price, has had the natural effect of spurring interest in uncovering new deposits, and in re-examination of known low-grade deposits and old tailing clumps. This is in addition to the usual reevaluation activities which normally result from the constant improvements in efficiency of ore concentration.

Jan 1, 1949

By William R. Yernberg

SME’s Arizona Conference was held at the Doubletree Hotel at Reid Park in Tucson on Dec. 4 and 5, 2005. The annual conference is organized by the Arizona Conference Board of Directors, which represents SME sections in Arizona and adjacent areas. The conference provides an opportunity for individuals and corporations associated with mining in the region to get together once each year to discuss the status of the industry. Important topics discussed among the attendees included the price of copper, community relations and education.

Jan 1, 2006

By John A. Wolfe

The brightest spot in the mineral economy of the Philippines is the rapid development of reserves and production of copper. If economic conditions remain favorable and the political climate is encouraging, production of this metal should increase at a rate of 25% per year for several years, long enough to rank this country as one of the major copper producers of the world. In 1965 the copper content of the Philippine concentrates was approximately 60,000 tons. It should total about 80,000 tons in 1966. A major part of the increase will come from the rich Lepanto mine located in Mountain Province, North Central Luzon. Daily production of ore increased from 1200 to 2000 tons at the end of 1965. The ore, consisting of enargite and luzonite, averages nearly 3% copper and 0.12 oz. of gold per ton. It is an under- ground operation with a cutoff grade of 2%. The ore is associated with extensive silicification under a volcanic cap rock. Further expansion to increase ore production to 2600 tpd is under way.

Jan 4, 1967

By R. L. Player

"The Copper Isasmelt process has been developed over the past 15 years by Mount Isa Mines Limited (MIM) and the Commonwealth Scientific and Industrial Research Organisation (CSIRO). The first two commercial plants were commissioned in mid 1992.The results of four heat balances on the Mount Isa plant are presented. At a smelting rate of 98 t/hr of chalcopyrite concentrate plus 14 t/hr reverts, the total heat generated was -91MW. Approximately 39% of this energy came from coal, while 50% of the energy reported to the process waste gas. To a first approximation the furnace is near chemical equilibrium.The impurity distribution in the Mount Isa plant was presented. Approximately 91 % As, 85% Cd, 75% Bi, 68% Zn, 60% Sb and Tl, 45% Pb and 30% Te in concentrate was eliminated from the matte.The ""bubble"" frequency of the gas cavity below the foot of the lance was measured at -2. 1Hz. Side bands displaced by -l.lHz modulated the bubble frequency. This was consistent with radial gravity wave motion in the slag bath. The two processes, gas bubbling and wave motion appeared to be coupled together; the bubble frequency was twice the slop wave frequency. IntroductionCopper Isasmelt has been developed over the past 15 years by Mount Isa Mines Limited and the Commonwealth Scientific and Industrial Research Organisation. The process is based on the Sirosmelt lance, invented by researchers at the then Division of Mineral"" and ProcessEngineering ofCSIRO in 1973.The first two commercial scale copper furnaces were commissioned at Mount Isa and the Cyprus Miami Mining Corporation smelter at Miami, Arizona in August and June 1992 respectively.The process concept, the history of the development, and the Mount Isa plant have been described (1, 2, 3).This paper sununarises heat balances obtained on the furnace, the impurity distribution among the output streams and lance injection characteristics."

Jan 1, 1996

By G D. Richmond, J C. Chomley

Copper has been mined intermittently at Gunpowder, north of Mount Isa, since 1927. Most ventures have had some initial success but fluctuating copper prices, the small-scale of the operations and the remoteness of the site led to closures or sale. Processing routes have included direct sale of high grade ore, flotation of oxide and sulphide concentrate, and copper cement production from mine solutions. Copper cathode has been produced by heap and in-place leaching coupled with solvent extraction and electrowinning (SX/EW). Aberfoyle Limited recently purchased the Gunpowder operation and is expanding operations. An innovative high recovery tank leaching process, including a low temperature, low pressure autoclave, has been adopted. Copper cathode will be produced from an expanded solvent extraction/electrowinning plant at approximately 45 000 tonnes per year. The combination of expanded throughput and high metal recoveries will make Gunpowder a cost competitive producer.

Jan 1, 1998

By Jaeheon Lee, Junmo Ahn, Jiajia Wu

In this project, low grade copper ore samples were obtained from two mining sites in the US, and they were chalcopyrite-rich (sample A) and conichalcite-rich (sample B) ore. Copper grades were 0.6% and 0.4% for the two samples, A and B, respectively. Leaching experiments were conducted by bottle roll tests for 24 hours, and various lixiviant systems were applied. Methanesulfonic acid (MSA) and glycine were used as alternative lixiviants to sulfuric acid. Oxidants including ferric ion and hydrogen peroxide were tested. The results showed the copper extraction of sample A increased after adding 1 g/L ferric ion in sulfuric acid and MSA solution, from 20% to 36%, but further increase was not observed with higher ferric ion concentration. The copper recovery increased significantly with H2O2 concentration, and over 80% copper recovery was obtained with 10% H2O2. For sample B, the copper dissolution kinetics were relatively fast and reached the maximum value in five hours. Around 60% copper were extracted in sulfuric acid and MSA solution. Copper recovery slightly increased after adding ferric ion and H2O2. Glycine is a weaker lixiviant for both samples, the copper extractions were 36% and 4% respectively. INTRODUCTION Chalcopyrite is the most abundant copper sulfide mineral on earth, and it is one of the most refractory copper minerals regarding chemical processing. The declining copper grade and depletion of copper oxide deposits stimulate the development of hydrometallurgical process to extract copper from chalcopyrite. The dissolution of chalcopyrite by ferric ion is normally illustrated as Equation (1) and (2) (Dutrizac & MacDonald, 1974): CuFeS2 + 4Fe3+ → Cu2+ + 5Fe2+ + 2S0 (1) CuFeS2 + 4Fe3+ + 3O2 + 2H2O → Cu2+ + 5Fe2+ + 2H2SO4 (2) However, slow kinetics are usually observed from chalcopyrite leaching at ambient temperature and pressure. Researchers attribute the slow kinetics to passivation during chalcopyrite dissolution, and different passivation products are also proposed (Klauber, 2008). In recent years, many efforts have made to develop new methods to extract copper from chalcopyrite, such as utilizing bacteria or stronger oxidant (Zhao et al., 2015; Ruiz-Sánchez & Lapidus, 2017).

Jan 1, 2019

By P. Blazy

A hydrometallurgical procedure for the recovery of metal from ore, particularly of copper from sulphide materials is described. In the course of this procedure, the raw material is first of all submitted to the action of an appropriate reducing system and then treated with a ferric salt solution which brings about the selective solubilisation of the elements to be recovered with respect to those constituting the pyritic phase. This procedure is investigated using two ore samples proceeding from massive sulphide deposits, presenting different chemical and mine- ralogical characteristics. The results of this investigation are compared with those obtained by a classical process employing a ferric salt solution. It appears that one solubilisation of copper takes place under particularly advantageous conditions when our procedure is used. The principal parameters affecting leaching of zinc, lead and copper are examined. The selectivity of this treatment is evaluated by following the change during the treatment in the pyritic phase.

Jan 1, 1976

By G. Sedelnikova

This paper is aimed to elucidate the problem of copper losses in pyrometallurgical slag. Currently, flotation is used to recover copper from the slag and return the metal into the pyrometallurgical process, but the percentage of recovery frequently remains unacceptable. The process of copper leaching from the slag flotation tailings using biologically regenerated ferric sulfate is considered here as an approach to the issue. A sample of the slag flotation tailings containing 0.67% Cu and 4.99% Zn was supplied by the Urals Mining and Metallurgical Company (UGMK, Russia). The Cu and Zn speciation analysis of the sample demonstrated that Cu occurred as 58.1% oxide, 15.4% native Cu, and 18.1% secondary sulfides whereas chalcopyrite proportion was as low as 8.4%. Prevailing Zn species were refractory (58.9% Zn ferrite), and zinc oxide and silicate species made just 37.2% of the total metal content. The silicate generated silica gel that strongly hampered filtration. As demonstrated, Cu recovery is a possibility, which is apparently not the case with Zn. In this work, the stirred tank bioreactor acid leaching employing bacterial Fe3+ sulfate solutions was studied, with ferric sulfate being a product of Fe2+ biological oxidation and not the chemical reactant adding of Fe3+ sulfate. Acidithiobacillus, Sulfobacillus, and Leptospirillum bacilli were employed in Fe2+ biooxidation. Temperature, pH, Eh, S:L ratio, leaching time, and Fe3+ concentration in the source leachate as the process controls were considered. As demonstrated, the process employing biologically regenerated ferric sulfate proved to be more efficient than conventional chemical leaching. Under the optimum conditions, the proportion of Cu extracted into leachate during 2 hours was 79.1%, and that of Zn 27.2%. Precipitation of high-grade Cu sulfide served to extract the metals from the leachate, and Cu extraction was 99.7%. All in all, 23.2% Cu, 0.03% Zn, and 0.02% Fe was returned into the pyrometallurgical process. Keywords: slag flotation tailings, copper leaching

Sep 1, 2012

By E. E. Malouf

For centuries small amounts of copper have been recovered from acid mine-drainage waters. In recent years, the expansion of copper-bearing waste dump leaching for copper recovery has established a technological basis that is rapidly converting the art of copper leaching' into a science. Leaching of copper, whether from copper-bearing ore or waste, follows well established chemical reactions. These reactions are commonly in an acid circuit, but where leaching of copper is to be made in rock that is a high acid consumer, ammonia and ammonia carbonate, as well as cyanide leach solutions, can be used effectively. Current hydrometallurgical processes are being applied at an ever increasing rate to recover the trace amounts of copper in copper mine waste by in-place leaching. Currently approximately 300, 000 tons of copper are produced annually in this manner where as recently as ten years ago, only one-eighth of this amount of precipitate copper was produced.. Where copper occurs as an oxide in higher concentrations as to be classified as an ore rather than waste, then the hydro-metallurgical processes are applied as percolation or agitation leaching in vats, pachuca tanks, or prepared heaps. This approach gives better control over the recovery of the leach solutions, particularly in the processes where ammonia or cyanide leaching agents are to be used.

Jan 1, 1972

By R. Scheffel, R. Washnock, G. Zarate

"The second survey of global copper leaching operations has been completed, the data published in 1998 has been updated. A survey questionnaire was sent to 40 identified copper leach operations around the world including those leach properties still in operation that responded to the 1998 survey. This paper lists the results of 26 operations that responded to the survey. Selected information from the survey is represented in graphical form. INTRODUCTION Copper leaching was first performed by truck dumping huge monolithic stockpiles of ore and flooding the surface with leach solution or lixiviant with little understanding to the parameters that support optimum recovery of the metal. Heap leaching has evolved into a controlled process whereby a complex low grade ore is stacked in shallow lifts (under 10 m in most cases). The ore is usually crushed and agglomerated, conveyor stacked on an engineered containment system (leach pad) and irrigated in a controlled manner. The lixiviant chemistry is designed to extract the optimum amount of metal from the ore. At first glance, heap leaching appears to be simple in nature. In fact, it is a very complex process. A plethora of technologies such as material handling, environmental control, analytical techniques, solution chemistry, hydrology and many others contribute to the design and operation of a heap leach. Heap leaching requires an inter-disciplinary approach between geologists, mining engineers and metallurgists. Many operational factors affect heap leaching. The process must address climatic factors, flexible operation, quality control before and after leaching, physical characteristics of the ore, geology, mineralogy, reagent consumption, solution application rate and methods (wobblers or emitters), particle size, and favorable conditions for bacterial growth and survival. Predicting, troubleshooting, and optimization of leach recovery is a significant challenge when considering the number of variables involved and their interactions The survey updates many of these key operating parameters to the current state of the art. It should provide the reader an excellent benchmarking summary of these factors on a site-by-site basis."

Jan 1, 2016

By Ivan Imris

The common problem of new and conventional technology of copper production is the copper losses in the slags, which may be divided into mechanical and physico-chemical losses. In the first group can be included particles of mechanically entrained or floated unsedimented matte particles which coexist with slag. On the other hand, the physico-chemical losses are caused by solubility of copper in the slag in the form of sulphide and oxide. The copper losses in the slags from different processes has been predicted by calculation from thermodynamic data and compared with those determined by microscopic examination and quantitative electron microprobe analysis. Depending on the forms of copper losses in the slag the reduction of copper losses in the slag or the slag cleaning process could be suggested.

Jan 1, 2003

By R. W. Ruddle

The present state of knowledge on this subject may briefly be summarized as follows: (1) The copper content of the slag increases with the grade of the matte in contact with it; on the other hand, the proportion of copper lost decreases as the matte grade increases, because as a general rule a smaller volume of slag is produced when a high-grade matte is made. (2) The presence of CaO and MgO in the slag decreases the copper content of the slag, and the presence of ZnO increases it; the presence of A1203 seems to have little effect. The influence of the silicate degree is not yet clarified, but on theoretical grounds it seems likely that the copper content increases with the silicate degree of the slag. (3) High slag viscosity, high magnetite contents and low density difference between matte and slag all tend to increase the copper loss. The volume of slag is also an important factor. (4) The effect of bath temperature on the copper loss is not clear and further investigation of this point is needed. (5) It appears possible to recover a large part of the copper in the slag by adding FeS, CaS and other sulphides

Jan 1, 1953

By ?. ?ivkovic

The silicate slag and copper matte that are the byproducts of smelting copper in a reverberatory furnace contain SiO2, FeO, Fe3O4, CaO and Al2O3, and Cu, Fe and S, respectively. These components influence the oxide (Cuox), sulfide (Cusul) and total (Cutotal) copper content lost in the waste slag. This paper uses multiple linear regression analysis (MLRA) to determine the degree of influence. The results of calculations obtained using MLRA were compared to actual measured results and found to have a large degree of correspondence, with the coefficient of determination R2 = 0.975. These results indicate that slag composition and copper content in the matte influence the copper losses in the waste slag with a probability of above 97%. It was determined that, of the total copper content of the slag, 74% is in sulfide and 26% in oxide form, and that this ratio is not dependent on the slag or matte composition.

Jan 1, 2010

By M. Sanchez, I. Imris, G. Achurra

The Smelting and Converting El Teniente Process for copper concentrates as well as Slag Cleaning Furnace Technologies to treat slags coming out from the El Teniente Furnaces are widely applied today in Chile and around the world. Several industrial applications of this process have shown their capability to treat copper concentrates in a wide range of chemical and mineralogical compositions. Main operational parameters determining performance of the process are: oxygen enriched air flow rate, degree of oxygen enrichment, moisture content of the solid materials processed, molten material levels inside the vessel, frequency of molten materials tapping, bath temperature and copper losses in slags. In this work, the copper losses in the slags from El Teniente pyrometallurgical process are predicted by calculation from thermodynamic data and compared with those determined by microscopic examination and slag industrial data. Results obtained show that the main part of copper losses in the slag from El Teniente Furnace are mechanically entrapped or floated unsettled droplets of the coexisting matte phase. by the contrary the very low copper losses in the slag from slag cleaning furnace are limited by the physico-chemical form.

Jan 1, 2004

By David B. George

Kennecott Utah Copper adopted Outokumpu Flash Smelting and Kennecott-Outokumpu Flash Converting in 1995, a process based on granulating high grade copper matte. While nickel mattes and very small quantities of copper matte had been commercially granulated for years, no large scale copper matte granulation system had been built. This paper summarizes the history and benefits of copper matte granulation, the Kennecott design basis, the successful operation for over 10 years and potential problems inherent in some matte granulation approaches.

Jan 1, 2006

By H. M. Shepard

THE copper industry operated at high capacity throughout 1947, with no serious tie-ups in operation as was the case in 1946, when almost the entire industry was shut down by a four-month strike. Refined copper production in the United States averaged 94,250 tons for the first eleven months and deliveries to domestic consumers, including foreign copper for domestic consumption, averaged 114,000 tons per month. On April 29 the import duty of 4d per lb was suspended until March 31, 1949. Copper prices fluctuated considerably for a short time thereafter. Over the year the price of domestic electrolytic copper at refineries on the Atlantic seaboard advanced from 19.225¢ on Jan. 2 to 21.200d on Dec. 24.

Jan 1, 1948

As a result of the rapid economic growth in Japan, environmental problems have become a national concern in the latter half of the In 1967, the Basic Lair for Environmental Pollution Control was enacted as a fundamental ::olicy in this field. This lair provided compre- hensive guidelines for the action to be taken pollution prevention. The biggest problem n air pollution of those days was that of the sulphur dioxide content of the atmosphere. Based upon the law, the smelters in Japan have struggled hard to solve this difficult problem. Consequently, the ambient SO2 level has been lowered gradually year by year. At present, the ambient SO2 level is nearly equal to that in the atmospheric background. Furthermore, the fixation ratio of sulphur in Japanese copper smelters has attained up to 99.5 per cent.

Jan 1, 1980

By A. E. Isaacson

The U.S. Bureau of Mines is studying the effect of supercritical water, at 400' C and 3,500 psi, and subcritical water, at lower temperatures and pressures, on mixtures of minerals and metals to determine the possible applications of water at high temperature and pressure for processing these materials. This research was conducted using a flow-through bench-scale system. Copper-bearing minerals studied were chalcocite, chalcopyrite, covellite, and cuprite. In general, mixtures of copper minerals and pyrite reacted to form bornite. Intermediate phases formed were chalcopyrite and digenite. Similar reactions occurred between pyrite and metallic copper. A mixture of chalcopyrite and pyrite was inert in supercritical water'. The reaction between cuprite and metallic zinc yielded metallic copper. The cuprite was reduced by hydrogen generated by the reaction between metallic zinc and water. None of the reactions occurred exclusively in supercritical water, but also occurred in water at high temperature and pressure below the critical point. 'Experimental results did show that a dual-reaction process, consisting of solid-state diffusion and fluid transport, was occurring.

Jan 1, 1989