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By C. Cano

One of the main concerns within the mining industry is to minimize the impact of tailing facilities to the environment and nearby communities. Since the implementation of the first thickened tailings disposal (TTD) methodology at Kidd Creek in 1964, the thickened tailing disposal has been growing alternative disposal technology to mitigate the environmental impacts. Technical investigations on the TTD technology have been studied in detail to counteract the increasing shortage of water resources. One of the benefits observed with TTD is its resistance to potential earthquakes, particularly in zones with high seismic activity (Los Andes). TTD technology optimizes the maximum possible density for the material and generates a negative internal pore pressure in the material due to the surface and sub-surface water evaporation. This internal pore pressure should be enough to resist the pores pressure to which tailings are subject due to an earthquake action. The objective of this study is to verify the Shrinking Limit (SL) of thickened tailings during initial deposition and in the long term and the ability of tailings deposited by TTD would resist liquefaction. Fresh tailings coming from a flotation process of a copper plant concentrator were used. The samples put under cyclical three-dimensional tests associated with maximum credible earthquake magnitude range of 7.0 8.0 to develop a curve for the safety factor v/s depth for dry densities related to the maximum density obtained at the SL. At a dry density of 1.65 t/m3 liquefaction did not occur at any depth. At 90 % of dry density of 1.51 t/m3, liquefaction did not occur at earthquake magnitudes of 7.0 or 7.5 nor at any depth but liquefaction occurred at magnitude 8.0 between 0 m and 14.5 m; below 14.5 m, liquefaction does not occur.

Jan 1, 2007

By Doug Smith, Gary Taylor, Paul Ogryzlo

"Uranium mining and processing typically requires large quantities of high-grade reactive quicklime for the neutralization and clean-up of contaminated mine water and mill process waste streams. Currently the Saskatchewan uranium industry’s lime requirements are being supplied from deposits located in Manitoba and Alberta, resulting in transportation costs, which equal or exceed the market value of the product. During the summer of 2001, the northern village of Pinehouse located 220 kilometers south of the Key Lake uranium mill, staked two previously discovered limestone deposits. The relatively shallow deposits, known as Missi Island and Sleeping Giant are located in close proximity to one another, approximately 35 km south of Pinehouse adjacent to highway 914. This paper describes the results of a scoping study carried out by AMEC for the development of the two deposits based on a seasonal limestone mine and a continuous 40,000 to 60,000 tpy wood fired calcining plant designed to supply all of the quicklime needs of the Saskatchewan based uranium industry. Also described are the results of calcining and customer acceptance testwork on samples of lime prepared from recent Pinehouse limestone drill core. The potential benefits of the project include a significant new source of revenue for the community of Pinehouse, lower lime costs for the Saskatchewan uranium industry and employment for northerners in the mining, processing and wood harvesting operations.IntroductionInterest in the Pinehouse/Massinahigan River area for limestone was precipitated by the results of a 1977 industrial minerals drill hole program by the Saskatchewan Department of Mineral Resources. These results led to the acquisition of land in the area by Missi Island Mines Ltd who carried out extensive exploration during the period 1978-1982. They discovered a deposit on the southwest side of the Massinihigan River thought to contain approximately seven million tonnes of limestone with an average grade of 94.5% CaCO3. Missi Island carried out preliminary feasibility studies and prepared an Environmental Overview Statement. However, in spite of positive recommendations, the project was dropped in the mid 1980’s."

Jan 1, 2003

By Douglas C. Shumway

Mining in Cushenbury Canyon began in 1860 when John Cushenbury left his Holcomb Valley gold prospect and discovered a silver vein in the limestones of the present day Cushenbury Mine. Cushenbury Canyon is located along the western boundary of the Mojave Desert in southern California and on the north slope of the San Bernardino Mountains. The Holcomb Valley, immediately south of Cushenbury Canyon, was the most prolific gold producer in southern California. Gold mining dominated the area until World War II. Interest then turned to the Lucerne Valley limestone deposits when Kaiser Steel built a plant in Fontana, CA. The steel making process required large amounts of metallurgical grade limestone. In 1956, Kaiser Cement constructed a cement plant and limestone mine in Cushenbury Canyon. In 1982, Kaiser modernized the facility to a four-stage, preheater-precalciner system. Mitsubishi Cement Corp. purchased the property in April 1988. The company further modified the facility to create the most efficient and modem cement plant in the US.

Jan 1, 1992

By Marcelle A. Hauseux

"The Turgeon Lake granite lies 5 miles north of Baie Johan Beetz, Quebec, at the eastern margin of the Grenville Province. The rocks are classed as granites, granulites, quartzites, biotite schists and gneisses, and amphibolites. The granite's uraniferous character is manifested by phosphuranylite, uraninite, and metamict .zircons associated with magnetite. Field anomalies occur in irregular pods, conformable with the tectonic foliation, in bands roughly paralleling sedimentary bedding. Associated features diagnostic of uraniferous granite are smoky quartz, brick-red feldspar, magnetite and a coarse to pegmatitic grain size. The granite is chemically and physically heterogeneous. Metasediment-granite contacts are diffuse or sharp, and usually conformable to bedding. The regional foliations within all rock types are coherent, and sedimentary structures are often preserved within both metasediments and granite. Autometasomatism probably was a strong granitization agent.A uranium reconcentration into fold-axial zones from original sediments which have been highly granitized during at least two phases of folding is suggested from the available data."

Jan 1, 1977

IN CHAPTER I it was shown that the approximate market value of the copper produced at the twelve big Porphyry mines from 1905 to 1931, inclusive, was $2,821,300,000. This represented the refined metal ready for shipment to the consumer. In addition, at least $50,000,000 worth of gold and silver had been produced, more than $30,000,000 coming from the Utah mine alone. Consequently, the revenues totaled about $2,871,300,000, as shown in Table 58. Society paid this sum for a given quantity of metal that it wanted for industrial use. Where did the money go? Who were the direct beneficiaries of the creation of this new wealth? These questions this chapter will answer in a general way. Because some necessary data are not available, an exact tabulation in detail can not be made, but Table 58 is believed to reflect with reasonable accuracy the disposition made of the aggregate income from the twelve operations. The word "operations" is used advisedly because the complexity of corporate relations precludes the possibility of picking out companies, as could be done if each enterprise were an independent entity such as Miami, for example. Some explanation of the table and some elaboration of the items under "Outgo" is desirable. In the first place, it is assumed that the sum of surplus current assets existing in various treasuries at the end of 1931 was substantially equivalent to the “new” money that actually was put into the enterprises as a group. Less than $20,000,000 in new capital was realized from securities sold by the Utah Copper Co., the remaining funds coming from earnings. In respect to some of the others, such as Chuquicamata, Inspiration, and Andes, the amount of new money was much greater. However, of the two items-total new capital invested and total current surpluses-the former

Jan 1, 1933

By Christie AB, Bull VH

The geological environments of gold deposits are reviewed in relation to the strong resurgence in exploration for gold since 1979. The resurgence is due mainly to New Zealand's geological endowment with gold and its history as a gold producer. Major past production was from Cenozoic placers of Otago and Westland, late Cenozoic epithermal deposits of the Hauraki Goldfield, and the respective Paleozoic and Mesozoic lode deposits of the Reefton and Otago Goldfields. These same geological environments are the targets of current exploration, mainly in extensions of previously mined deposits. On the Westland and Otago Goldfields alluvial and fluvioglacial gravel placers, mainly representing lateral and depth extensions of previously worked ground, are being mined and prospected. Gold production in 1984 was exclusively from these deposits, and the development of mobile concentration plants has resulted in a great increase in mining activity. Mineralised zones adjacent to previously mined lodes in the Hauraki Goldfield, Reefton Goldfield, and at Macraes Flat (Otago Goldfield) are being intensively explored. At Waihi (Hauraki), an open-pittable resource of about 10 million tonnes of ore has recently been delineated. Some new deposit types and little explored environments are also receiving attention. They include: gold-sulphide quartz veins hosted in an alkaline granite dike in west Nelson; hot spring type gold deposits in fossil geothermal (epithermal) systems of Northland, the Hauraki Goldfield and the Taupo Volcanic Zone; and marine placers off the West Coast where an area of auriferous gravels has been outlined near Hokitika. There is potential in known deposits and prospects currently being evaluated for a large increase in gold production by the early 1990s.

Jan 1, 1986

By S G. Keyte

Gravity methods of recovering alluvial gold are not very efficient at particle sizes below about 0.l mm. Whereas such gold may not have been economic to recover in the past, the same may not be true today. Centrifugal concentrators, like the Knudsen and Knelson bowls, are showing promise after an inauspicious start, at least in New Zealand. The large proportion of heavy mineral in many auriferous beach sands and the natural hydrophobicity of fine gold should make flotation an attractive contender for its recovery. The conventional mechanical flotation cell is effective, but roughing, scavenging and cleaning are necessary to achieve both high recovery and low mass Yield. More problematic is the difficulty of maintaining coarse and dense sands in suspension. The air sparged hydrocyclone is effective at recovering fine gold from a fine Colarado River sand, but is not effective for the coarse and dense beach sands of New Zealand. The flotation column shows much promise for these sands, when recoveries of over 80 per cent and mass yields of less than one per cent can be achieved in a single pass. The froth separator of the USSR and flash flotation could also be worthy contenders. It is felt that flotation could be limited to naturally and artificially de-slimed concentrates.

Jan 1, 1990

By G. H. Clevenger

Discussion of the paper of ALLAN J. CLARK presented at the San Francisco meeting, September 1915, and printed in Bulletin No. 103, July, 1905, pp. 1381 to 1398. G. H. CLEVENGER, Palo Alto, Cal.-Since the precipitation of extremely dilute solutions as regards cyanide presents one of the greatest problems of zinc-dust precipitation, I should like to ask Mr. Clark the minimum concentration in cyanide of the solution from which it is possible to precipitate satisfactorily the gold with zinc dust in Homestake practice. A. J. CLARK.-For a day's work about 0.015 per cent. That is precipitating a very low-grade solution, very low in alkalinity. It requires probably 0.2 lb. of zinc dust to the ton. JOHN A. FULTON, Melones, Cal.-At the Melones mine, we have been working very low-grade rock for a number of years, the process being amalgamation and concentration up to the last year. The value of the ore increased so that we had to supplement the concentration with a sand and slime plant. Classification was a problem, since to install classifiers would mean tearing the mill all to pieces. We are therefore making use of a Wilfley table for this purpose. The amount of wash water is increased and a stream of fresh water is also run into the feed box at the

Jan 12, 1915

By K. Biswas, S. Iverson, S. J. Jung

Computer modeling of rock flow in ore passes is being investigated and compared using MSC Soft-ware’s Working Model 2D1 (WM2D) and Itasca Corp.’s Particle Flow Code in Two Dimensions (PFC2D). This work is being carried out by the Spokane Research Laboratory of the National Institute for Occupational Safety and Health to determine the usefulness of computer modeling to improve ore pass design. The types of tests included pendulum tests to validate WM2D simulations and simulated single-rock drop tests in both WM2D and PFC2D to understand their contact models. Simulated rock impacts at the ore pass chute and the use of inclined ore passes and doglegs were investigated in both programs. A PFC2D simulation showed that dynamic loads at the chute gate were reduced significantly when a dogleg transition was used, and simulations in which the ore pass was inclined showed significant reductions in chute impact loads compared to simulations of a vertical ore pass. A released hang-up was simulated using PFC2D.

By Milind Deo, Neale R. Neelameggham

The mixture of carbon dioxide and water vapor - called Soda [as in carbonated water] can be converted to easily stored transportation fuels. Earlier papers on Soda- fuel metallurgy noted that this conversion is endothermic requiring energy. Reduced state metallic elements - preferably the alkali, alkaline earth elements, or aluminum can act as oxygen carriers. Availability of such a combination in commonly land fille waste materials is shown. Plant based wastes which are normally composted form a better-controlled reagent. The energy required for the reaction can come from solar energy. This paper describes possible reactor designs, which can accommodate this conversion effectively using carbon dioxide recovered from flue gases.

Jan 1, 2008

By Milance Mitovski, Milorad Cirkovic, Vlastimir Trujic, Dragana Živkovic

"In this article, it is analysed the possibilitis and the effects utility waste energy of process sulfuric acid production in the existing technical-technological conditions. The utility second process energy, reduction the fuels consumption, reduction environment pollution and its effects are shown, too. High heat quantity is produced as the result of exothermal reactions in the production process of sulfuric acid in Bor. A part of this heat is sufficient for the process development, and significant part of this heat presents low-temperature heat energy that is irretrievable lost. Investigations were carried out based on production of 485 t of monohydrate/day of sulfuric acid. By recovery of this heat quantity for heating purposes, about 20.000 – 25.000 t of coal/year would be saved, what would significantly improve the power efficiency of the smelter copper production. IntroductionTechnological process of copper production in RTB BOR (Serbia) produces a significant quantities of high temperature secondary (waste) heat energy that is uncontrolled “lost” in the vicinity. Secondary heat energy is produced in pyrometallurgical copper production in quantity of 1,621 kWth/kg of dry concentrate or existing heat (that presents a difference of heat at furnace outlet and heat at temperature over temperature of dew point of sulphuric acid gases) in quantity of 1,315 kWth/kg of dry concentrate. For designed capacity of Copper Smelter in Bor, the existing heat power of secondary energy is 92,111 MWt, what is an equivalent to the power of steam-turbo plant where the Rankine-Clausius heating cycle is done, and that is 28,334 MWe."

Jan 1, 2008

By M. J. Ackerman, C. R. Waine, W. M. Merritts, L. P. Mokwa

Tests were conducted in the Pocahontas No. 4 coalbed in southern West Virginia to determine if successful degasification techniques, developed during an earlier study in the Pittsburgh coalbed, would be effective in removing methane from those usually more gaseous coals. Specially designed long-hole drilling equipment and Bureau of Mines continuous recording devices were used to measure the quantity and methane content of the ventilating air during the investigation. Infusions were made with water at normal waterline pressures ranging from 200 to 220 psi or at pump pressures ranging from 400 to 650 psi. Results of the tests showed that methane emissions from free-flow bleeder holes were as much as 131 cfm. Infusions of some of the holes increased the methane emission from other holes and from the exposed coal surfaces near the working areas as much as 1,540 cfm. The methane content in the main return-air currents after infusion was reduced more than 86 percent of the amount recorded before tests were made.

Jan 1, 1963

By Enrique Monge

Para resolver la crisis de combustible por la que atraviesa el Perú, se requiere reemplazar gradualmente el petróleo de consumo nacional por carbón mineral de los yacimientos del Santa, Oyón o Huayday. En ese sentido, el presente texto señala el tratamiento adecuado para la explotación del carbón a través de conocer el comportamiento de los carbones en los procesos de lavado, pulverización, briquetaje combustión, coquificación y gasificación. Para efectuar dichos estudios se recomienda la creación de un Instituto de Investigación Experimental, financiado por el Estado y por la industria -petrolera y carbonera, que opere en forma independiente, resolviendo cada uno de los numerosos problemas que afectan el tratamiento de los carbones como la liberación y separación de granos; expansión de carbón; fusibilidad de cenizas; combustibilidad y reactividad; rendimiento de gases; temperaturas de combustión; coquificación y gasificación; presión y exceso de aire para combustión; plasticidad, aglutinación y aglomeración de carbón, impermeabilidad, coherencia y resistencia de briquetas; etc

Jul 1, 1959

By F.S. Norcross

Manganese has long been considered one of the United States' most important strategic raw materials. Its indispensabil-ity in steel manufacture makes it vital to the nation's industrial life. Coupled with this indispensability is the fact that domestic producers have not been able as yet to supply more than a small part of the nation's manganese needs; consequently the United States has had to depend upon such distant foreign sources as the U.S.S.R., British India, the African Gold Coast and Brazil for the major portion of its ore. The great bulk of the mineral is consumed in the form of ferromanganese, an alloy that averages 80 per cent manganese and 20 per cent mainly iron and carbon. Manganese is used as a deoxidizer and desulphurizer in steel manufacture. By combination with the residual oxygen and sulphur of the bath, it helps to produce a clean, sound metal. Approximately 14 lb. of manganese, it is estimated, is used in making one ton of steel. For standard-grade ferromanganese, the ore should have a minimum of about 48 per cent Mn and preferably more. Beneficiation methods have not yet been developed for making the immense low-grade deposits in the United States salable at prevailing prices. Dependence upon foreign sources thrust the United States into serious difficulties during the World War. At the start of hostilities in 1914, the nation was obtaining about four-fifths of its ore from the Eastern Hemisphere. Domestic production had been negligible, amounting to only 3000 to 7000 tons a year. As the war continued, shipments from the East, endangered by sea attack and blockade, dwindled rapidly. Imports from Brazil mounted. In this grave emergency, the United States, starting almost from scratch and under the stimulus of high prices, succeeded by 1918 in producing a peak of 3II,000 tons, about 35 per cent of its consumption that year. Since the war, the United States has depended upon foreign sources for more than nine-tenths of its high-grade ore. During 1936, 1937 and 1938, imports for consumption averaged 736,000 tons a year. During the same period, domestic shipments averaged 21,000 tons a year. Of the imports, the U.S.S.R. accounted for an annual average of 280,000 tons, the Gold Coast 208,000 tons, India 74,000 tons, and Brazil 72,500 tons. But India and Brazil, which in 1936 ranked third and fourth in the list of importers, were displaced in 1937 and 1938 by a new and much nearer source—Cuba. From less than I per cent of United States imports in 1931, the island republic increased its shipments to more than 27 per cent in 1938. Cuba's average for this three-year period was 97,000 tons, and its total for 1938 was 131,000 tons. History The story of the economic development of manganese in Cuba dates back to the Spanish-American War days. John Green-way, then one of Theodore Roosevelt's

Jan 1, 1943

By F. S. Norcross

Manganese has long been considered one of the United States' most important strategic raw materials. Its indispensabil-ity in steel manufacture makes it vital to the nation's industrial life. Coupled with this indispensability is the fact that domestic producers have not been able as yet to supply more than a small part of the nation's manganese needs; consequently the United States has had to depend upon such distant foreign sources as the U.S.S.R., British India, the African Gold Coast and Brazil for the major portion of its ore. The great bulk of the mineral is consumed in the form of ferromanganese, an alloy that averages 80 per cent manganese and 20 per cent mainly iron and carbon. Manganese is used as a deoxidizer and desulphurizer in steel manufacture. By combination with the residual oxygen and sulphur of the bath, it helps to produce a clean, sound metal. Approximately 14 lb. of manganese, it is estimated, is used in making one ton of steel. For standard-grade ferromanganese, the ore should have a minimum of about 48 per cent Mn and preferably more. Beneficiation methods have not yet been developed for making the immense low-grade deposits in the United States salable at prevailing prices. Dependence upon foreign sources thrust the United States into serious difficulties during the World War. At the start of hostilities in 1914, the nation was obtaining about four-fifths of its ore from the Eastern Hemisphere. Domestic production had been negligible, amounting to only 3000 to 7000 tons a year. As the war continued, shipments from the East, endangered by sea attack and blockade, dwindled rapidly. Imports from Brazil mounted. In this grave emergency, the United States, starting almost from scratch and under the stimulus of high prices, succeeded by 1918 in producing a peak of 3II,000 tons, about 35 per cent of its consumption that year. Since the war, the United States has depended upon foreign sources for more than nine-tenths of its high-grade ore. During 1936, 1937 and 1938, imports for consumption averaged 736,000 tons a year. During the same period, domestic shipments averaged 21,000 tons a year. Of the imports, the U.S.S.R. accounted for an annual average of 280,000 tons, the Gold Coast 208,000 tons, India 74,000 tons, and Brazil 72,500 tons. But India and Brazil, which in 1936 ranked third and fourth in the list of importers, were displaced in 1937 and 1938 by a new and much nearer source—Cuba. From less than I per cent of United States imports in 1931, the island republic increased its shipments to more than 27 per cent in 1938. Cuba's average for this three-year period was 97,000 tons, and its total for 1938 was 131,000 tons. History The story of the economic development of manganese in Cuba dates back to the Spanish-American War days. John Green-way, then one of Theodore Roosevelt's

Jan 1, 1943

By Bruce Mathema, Steven Huynh, Terry Cheng, Gerry Gomez

"The use of PM40N, a specialty precious metals promoter in conjunction with a standard Xanthate based collector was extensively investigated in the flotation of a complex, silver-bearing, lead-zinc ore containing naturally occurring graphitic gangue. The ore, considered difficult to treat with conventional reagents, was tested at plant scale to determine the impact of the promoter on silver recovery, lead grade and also on the amount of zinc misplaced to the final lead concentrate.Initial laboratory-scale rougher flotation testing implied significant improvements in the recovery of silver and lead with a marginal amount of zinc displaced. Subsequent mini pilot flotation tests further revealed that a gain in selectivity and recovery could be achieved by adjusting PM40N promoter dosage and incorporating a cleaner flotation step, further improving rejection of sphalerite. Plant trials resulted in an improvement of 3% in Silver recovery coupled with a 2% increase in recovered lead whilst maintaining lead grade in the lead circuit.To date, the use of PM40N as a promoter has enabled the processing operation to significantly increase its plant economics. The partial replacement of the standard Xanthate based collector has also reduced the overall quantity of Xanthate required in the standard circuit by approximately 10% - which is a positive impact on the environment. It is envisaged that further optimization of plant parameters will favor stable operation and continue to boost revenues generated by subsequent improvements in silver metallurgy."

Jan 1, 2014

By R. Hogg, D. W. Fuerstenau

Based on a simplified model for the motion of particulate materials in horizontal, rotating cylinders, approximate expressions have been derived relating the power consumption of a tumbling mill to its dimensions and speed of rotation and to the amount and nature of the charge. The basic assumptions underlying the model are that there is no slip between the charge and the mill shell, that the free surface of the charge remains essentially planar during rotation, and that there is no transfer of momentum from the material flowing down the free surface back to the mill shell. The relationships formulated from this model have been shown to be in excellent agreement with experimental results taken from the literature. While the requirements of the model are most clearly met at low speeds of rotation, the results appear to be valid even for speeds up to 70-80% of the critical speed. Poorer agreement between theory and experiment is found when the charge consists of relatively large particles (balls) and when the filling of the mill exceeds 50% of the available volume. Qualitative explanations of these deviations have been offered.

Jan 1, 1973

High-pressure grinding rolls (HPGR) are an accepted, but continuously expanding, comminution technology. The copper industry has a growing need to process harder ores with progressively lower grades, more challenging metallurgy and increasing demand for efficient bacterial leaching applications. This has resulted in a major focus on HPGR use in a crushing circuit. If the industry is to increase leach extraction while lowering energy costs in high-throughput hydrometallurgical operations, HPGR will be of inherent significance. The following paper summarizes test work on copper leach ores utilizing the benefits of HPGR comminution, while increasing several critical metallurgical parameters. Based on the current work and previous investigations, it was found that HPGR comminution within a crushing circuit, in specific instances, can result in 2 - >10% increases in cop-per extraction. The tested variables are crush size, rock type and copper mineralogy dependence. Various other process parameters may also be optimized.

Jan 1, 2011

By Eduardo Amado Marconato

Na usina siderúrgica de Cubatão da Usiminas, o forno de reaquecimento de placas do LTQ2 utilizava como gás combustível a composição de gases de coqueria, alto forno e natural. A partir da interrupção da produção das áreas primárias (Redução e Aciaria) em Cubatão, o equipamento precisou ser adaptado para uso exclusivo do gás natural como gás combustível. Além das adequações feitas nos sistemas de controle de combustão foram adotadas ações para redução do consumo de combustível como desenvolvimento de novos procedimentos de acendimento do forno, reduzindo o número de queimadores ligados visando aproveitar melhor o calor gerado pela combustão do gás natural. O sistema de combustão foi reconfigurado, permitindo reduzir a temperatura da zona de pré-aquecimento e o tempo de permanência no forno, com isso a economia alcançada no consumo térmico superior a 7%.

Oct 30, 2017

By J. Lee, M. Peart, T. Bhambhani, D. R. Nagaraj, P. Riccio

"Over the past decade or so, our efforts have focused on developing the rationale and concepts for new technology in reagent development and optimization in the processing of base and precious metals. A major outcome of these efforts have been the development of a novel family of collectors, principally driven by the MAXGOLD® and Aero XD-5002® chemistries, for precious metals recovery from primary Au ores and Cu-Au ores. These chemistries was developed to be robust to address the complex mineralogy and gold deportment in these ores. Plant trial data for a variety of ores have confirmed that the both Aero® XD-5002 and MAXGOLD® chemistry are indeed versatile in recovering gold values occurring as gold alloys and/or associated with many types of sulfides, and in altering the deportment of certain sulfides (such as those bearing As, Bi etc). An overview will be given in this paper using plant data to highlight the application of these reagents.INTRODUCTIONThe subject matter of this paper is the flotation of Au containing minerals from their ores. The motivation from this manuscript comes from the several challenges faced with the processing of Au ores, as was articulated in detail by Nagaraj (2008). The biggest challenge is the complex mineralogy; most easy-to-treat free-milling ores bodies are largely depleted. The Au values are distributed in multiple host minerals such as pyrite, pyrrhotite, arsenopyrite and Cu minerals such as chalcopyrite. Au values also happen to be finely disseminated within the various minerals and are sometimes invisible in optical microscopy. Additionally, on any given day, some mines treat multiple ore types, and the mineralogy is different for each ore type.Xanthates continue to be the collectors of choice at most Au and PGM operations around the world. Most primary Au operations tend to consume very large amounts of collectors; the dosage ranges anywhere between 40-400 g/t and this has been attributed to both mineralogical and non-mineralogical factors (non-mineralogical factors include the low cost of xanthates and the mindset that “more is always better”, which is not necessarily the case). In addition, several studies now show that xanthates alone are never sufficient where multiple value minerals are present, and that supplementary collectors are required. The large doses present a substantial cost to the minerals industry, and many mining companies have expressed the need for more efficient reagents that can improve metallurgical performance at low dosages."

Jan 1, 2013