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By Richard S. Kunter

The Bulldog CIP plant treats the slime portion of current tailings and reclaimed flotation tailings from the adjoining 350 tpd (3.67 kg/s) silver-lead flotation mill. Changes to the dredge reclaimer, pulp storage tank and trash removal system have allowed an increase in tonnage treated to a sustained 460tpd (4.83kg/s). CIP feed is 3-4 oz/T Ag (102.9-137.1 ppm). Following conventional cyanidation, carbon is added to the slurry. Silver loaded carbon is stripped in atmospheric-pressure vessels using hot caustic cyanide. Silver recovery from this solution is by Zadra- type cells. Some of the modifications to the plant and operating techniques are discussed.

Jan 1, 1981

By R. J. Reuss

The ore at this new underground operation is a massive sulfide breccia with high copper values. A major highway and electric transmission line pass directly over the near-surface orebodv. The mining method being used is undercut and fill. A continuous miner together with a single boom, hydraulic jumbo is used for the mining. Backfill is a weakly cemented sand-gravel mixture that is pneumatically placed. The ore is treated by conventional crushing, milling and flotation.

Jan 1, 1992

By Yi Luo

A massive roof fall occurred in a small room-and-pillar coal mine. The roof falls occurred under a barrier pillar in a closed coal mine located a short distance above the active mine. The extensive fractures connected the active mine with the old mine, turning them into a “combined mine.” Since completely sealing off the roof fall area is prohibitively expensive for the mine, ventilating the connected old mine became the only option. A comprehensive investigation was conducted to understand the atmospheric condition in the old mine and to find an efficient way to ventilate the “combined mine.” A set of studies involving geology, gas content, pillar stability and mine subsidence were performed to explore whether the closed mine is filled with an explosive air-methane mixture. The air quantity measurement data were used to determine the resistance of the air path formed through the fractured zone. Observations and measurements made from a tour to the underground and mine surface enabled the formulation of simple but effective ways to ventilate the old mine. The ventilation network analysis also suggested methods of control, so that sufficient ventilation air can be directed to the two parts of the old mine.

Jan 1, 2010

By Joseph W. Murray

GENERAL DESCRIPTION The undercut-and-fill method of mining was devel¬oped by Inco Ltd. in the Sudbury district of Ontario, Canada, to deal with abnormal ground conditions en¬countered in pillar recovery. Although it is now some¬times utilized as a primary stoping method, the main application continues to be in pillar recovery. In any mining method in which fill is employed, pillar systems tend to fail gradually as stoping pro¬gresses. As a result, most pillars are completely destressed and broken by the time pillar recovery com¬mences. Control of the back in such ground becomes a major problem and is contingent on local conditions as they exist in the workplace. In most mining methods, operations such as drilling, blasting, slushing, and tim¬bering can be readily assessed and the mining cycle pre¬dicted. However, in the mining of pillars with overhand methods, the time and effort required to maintain a safe back varies to a quite unpredictable extent. For years the square-set method was the most effective means available for pillar recovery. However, in areas where bad ground was encountered, square setting was a costly and sometimes hazardous method. As the tonnage mined from pillars increased, it was apparent that a safer and less expensive approach was required. Past experience indicated that the best way to deal with broken ground was to mine from the top down. Top slicing systems were examined but were rejected due to the major sub¬sidence which resulted. However, ideas generated from studying top slicing finally developed into the present method of undercut-and-fill. This method, which has almost completely replaced square-set mining, exhibits reduced costs, has improved safety and efficiency, elimi¬nates the factor of bad ground overhead, and allows a closer prediction of the actual mining cycle. The result is more accurate scheduling of pillar production. This factor is of increasing importance as the tonnage from pillar recovery increases in relation to total mine pro¬duction. Undercut-and-fill is a method of extracting a block of ore by mining successive cuts, working from the top down. After a cut of ore is completely mined out, lami¬nated timber stringers are constructed along the sides for the full length of the cut. Round logs covered by a wire screen are laid across the stringers to form a mat. The cut is then tightly filled with hydraulically placed cemented sand fill. Mining is then resumed on the cut immediately below the mat. As the cut advances, the timber stringers are supported by round timber posts seated on the solid bottom of the cut. Drilling, blasting, muck removal, and timbering operations are repeated until all the ore in the cut has been removed. At this time, another log mat is laid and the cut is filled. This sequence is continued until the block has been mined from level to level. The basic principles of undercut-and-fill can be adapted to a variety of mining situations. It can be applied transversely or longitudinally in ore bodies of varying width, that dip from vertical to flat lying. The method can be applied using conventional equipment or it can be mechanized through the introduction of small diesel or air-powered units for muck removal. Undercut¬and-fill is a selective method, ore recovery is high, and dilution can be controlled to an acceptable level. How¬ever, it is a high-cost, labor-intensive method and as such its use is restricted to areas where the less costly meth¬ods of ground support are not successful or where bulk mining methods are not adaptable. One basic require¬ment for undercut-and-fill is an adequate source of hydraulically placed cemented sand fill and a dependable sand plant and distribution system. The success of the undercut method is to a large degree dependent on the competence of the fill placed in each of the successive cuts. In theory, the weight of the fill on each cut is supported in part by the cemented sand itself as it binds to the rough sides of the mined¬out areas next to the pillar walls, and in part by the tim¬ber support. As mining progresses beneath the mat, a downward pressure is exerted within the cemented fill and a voussoir arch develops above the mat, across the full width of the cut. This supports the major portion of the load. The log mat and posts have a double func¬tion in that they carry the remaining weight not sup¬ported by the voussoir arch and in addition prevent any pronounced settling of the fill mass. Excessive settling breaks the bond between the cemented fill and the pillar walls, eliminating the arch effect and transferring the entire load to the mat and posts. The theory of a voussoir arch forming within the cemented sand fill mass is substantiated in practice by the fact that pillars have been mined from level to level with no more significant weight on the bottom cuts than was experienced on the top cuts. In effect, as mining progresses downward, there is no additional transfer of weight from previous cuts to the cut being mined. APPLICATION OF UNDERCUT-AND-FILL Because of their size, the Sudbury ore bodies are, for the most part, mined transversely and, to a considerable extent, by filling methods. Transverse stoping blocks are designed with 9-m (30-ft) crown pillars established be¬low each level and with stopes separated by vertical rib pillars. In past years, floor pillars were left above the level as well, but present practice favors the silling of primary stopes at the base of the rail. Rib pillars are designed to meet the ground control requirements of the primary stoping methods. Local conditions and past experience generally dictate their size. Historically, since the square-set method was to be used for recovery, rib pillars were designed to accommo¬date standard 1.7-m (5'h-ft) sets. A typical rib pillar is four sets wide [6.7 m (22 ft)] when located between two square-set or cut-and-fill stopes having a width of 8.4 m (27.5 ft). When the undercut-and-fill method was de¬veloped, a mat width of two sets was adopted. This di¬vided each cut of the common four-set pillar into two

Jan 1, 1982

When a company like Lonmin, the third largest platinum producer in the world, is mining one of the most precious metals in the world, time is money. Every piece of equipment is expected to be in top shape because every minute of downtime can add up to lost profit. Two different types of platinum ore are mined and processed by Lonmin, namely Merensky and UG2. Merensky has a grade of 3.5 g/t (0.1 oz/st) and UG2 has a grade of 4.5 g/t (0.13 oz/st). The mine also produces platinum group metals (PGM), which consist of platinum, osmium, palladium, rhodium, iridium and ruthenium. Although South Africa has more than 80 percent of the world?s PGM reserves, these minerals are very scarce. Lonmin is considered a low-cost producer of these minerals, generating 21.5 t (694,000 oz) of platinum and 18.6 t (600,000 oz) of PGMs in 2012. So when the company?s Karee K3 concentrator plant in Rustenburg, North West Province, South Africa, was experiencing high operating costs on the transfer points, a solution was needed.

Jan 1, 2012

Keeping an eye on the future while honoring the past, Frank McAllister, chief executive officer of Stillwater Mining Co., and life member of the National Mining Hall of Fame and Museum, gave the keynote address to a packed house during the 23rd National Mining Hall of Fame and Museum Banquet in Leadville, CO on Sept. 11. McAllister?s company made news just days before the induction ceremony, when it was announced that Stillwater Mining would acquire Marathon PGM Corp. for $118 million (ME, Oct. page 14).

Jan 1, 2010

By W. M. Aubrey

The use of solid fuel in shaft furnaces to indurate iron ore pellets was originally considered a necessary evil, which the industry worked long and hard to eliminate. Since then changes have taken place in furnace design and operating practices that have led us to our suggestion that solid fuel additions could again be used. In essence, what has happened is that the developments which made it possible to operate without solid fuel have also created new operating conditions favorable to the effective use of solid fuel. Because these new conditions* were brought about in part by the desire to eliminate solid fuel and because the development was a gradual one, the industry failed to reassess its position and for years neglected the possibility of returning to solid fuel.

Jan 1, 1967

By James R. Hindman

Vermiculite is the mineralogical name generally applied to a group of hydrated, trioctahedral ferromagnesian aluminum platty or micaceous silicate minerals. In microscopic particles the mineral as mineral is differentiated from montmorillonite by a higher Si:AI ratio of approximately 3:l and a higher layer charge and higher cation exchange capacities. Commercial vermiculite, the vermiculite which is mined, beneficiated, and discussed in this chapter, is an important industrial mineral which is formed by the alteration of macroscopic particles of biotite and iron-bearing phlogopite. In natural form it has the size and shape of mica, while possessing the ion exchangeable properties of zeolites and some clays. Within the internal structure of vermiculite lie water molecules, which when rapidly heated to high temperature, transform into steam causing the vermiculite particles to increase in volume. This process of thermal exfoliation produces a lightweight product which finds use in various construction products, agriculture, horticulture, and other industrial applications. In exfoliated form it has the low density and biological inertness of expanded perlite, while maintaining a chemically active surface. Most of the uses of vermiculite are for the thermally exfoliated material. In addition, the cation exchange potential of vermiculite and its layer-silicate structure are properties which can be used as the basis for numerous products including intumescent coatings and gaskets, inorganic films, and the treatment of nuclear and other toxic waste. When chemically and physically modified, it possesses some properties common to organoclays and synthetic compounds. In high purity concentrates and manufactured products, these properties can be altered and optimized in applications to produce superior performance. GEOLOGY Mineralogy Vermiculite was first described as a mineral by Thomas H. Webb in 1824. Webb believed vermiculite to be a variety of talc and described its diagnostic property of thermal exfoliation thusly: "If subjected to the flame of a blowpipe, or that of a common lamp, it expands and shoots out into a variety of fanciful forms, resembling most generally small worms having the vermicular motion exact." Observing that with one exception, all commercial vermiculite has formed by the alteration of iron-bearing phlogopite, a simple definition of vermiculite might be the alteration product of trioctahedral micas, macroscopic particles of which display marked exfoliation when heated rapidly to temperatures above 374°C (the critical temperature of water). The first vermiculite described was found near Worcester, MA, but it was later pointed out by Brush (1866) that vermiculite had been known for some time in Japan as a children's novelty. Because the techniques for mineral analysis were somewhat primitive com- pared to those routinely available today, and because the structure and crystal chemistry of the micas and related minerals were not widely understood, the early literature generated at least 20 "new" minerals which were nothing more than variations in the color and chemistry displayed by vermiculite. Trade names such as Dugarlite, Mandolitem, Strong-LiteTM, and Zonolitem are sometimes used to denote commercial vermiculite, thermally exfoliated vermiculite and their products. Vermiculite particles resemble mica and vary in color from colorless through tan and green to black, depending on the chemical composition. Most physical and optical properties are close to those of biotite, with perfect basal cleavage, hardness between 1.5 and 2.5 Mohs, and specific gravity between 2.2 to 2.8. Fine-sized particles will feel talcose, especially when wet. However, for commercial vermiculite the diagnostic test is that of thermal exfoliation. Vermiculite can display a wide range of compositions, depending on the composition of the original mica and the change of chemistry during weathering and ion exchange. Chemical data for some commercial vermiculites are given in [Table 1]. There are numerous analyses for vermiculite given in the literature, and the analyses of vermiculites from commercial deposits will fall within the range indicated by the average analysis in the table. Many analyses separate the adsorbed water (H2OP) from crystalline water and water as hydroxyl (H20+). The basis of separation of adsorbed water is generally considered to be 100°C. However, hydrate water within the vermiculite interlayer begins to be lost at 75°C and lower. Therefore, measured water content and loss on ignition are combined in the table as H20[total].

Jan 1, 1994

By M. A. Reyes, J. Bickson, J. R. Srednicki

Federal laws mandate the presence of refuge alternatives (RAs) in underground coal mines and provide guidelines for harmful gas removal. After an explosion, harmful gases can follow survivors into an RA and can reach a deadly concentration. Investigators from the National Institute for Occupational Safety and Health (NIOSH) tested 12 ventilation system configurations (VSCs) to purge sulfur hexafluoride (SF6) tracer gas from a built-in-place (BIP) RA in the Experimental Mine in Bruceton, PA. Using a borehole air supply, tests were conducted with airflows of 750 and 1,000 standard cubic feet per minute (SCFM). Excluding a VSC where the air inlet was 1.98 m (78 in) from the air outlet, the results show that the time to purge the SF6 concentration from 1,000 parts per million (ppm) to 25 ppm ranged from 26 to 29 minutes for an airflow of 750 SCFM and ranged from 18 to 23 minutes for an airflow 1,000 SCFM. Mines can use this information to reduce purge time in case of a mine disaster and determine which VSC to avoid. INTRODUCTION After a fire or an explosion in an underground mine, the concentration of carbon monoxide can reach a fatal level. In a mine disaster, the main goal is to escape. However, if miners cannot escape, they have to migrate to a location to find refuge and, if possible, communication, and a borehole source for a fresh air supply. To ensure survival of the miners, Federal laws mandate the presence of a refuge alternative (RA) within 304.5 m (1,000 ft) of the nearest mining face. In addition, Code of Federal Regulations Title 30 CFR Section 7 sets the guidelines for personnel tracking, RAs, and the removal of noxious gases in underground coal mines (Mineral Resources, 30 CFR, 2014). Federal regulations define an RA as a protected, secure space with an isolated atmosphere and integrated components (Mineral Resources, 30 CFR, 2014). There are three types of RAs: portable metal RAs, portable tent-type RAs, and built-in-place (BIP) RAs. A BIP RA is constructed using masonry or other stoppings to create an isolated space in a mined-out area. Blast doors are installed in one or both of the stoppings for the miners to enter the RA. Furthermore, RAs must provide the following components: breathable air, food, water, air monitoring, and harmful gas removal. Federal regulations have emphasized the mandate for breathable air and harmful gas removal in the following words: “Purging or other effective procedures shall be provided for the airlock to dilute the carbon monoxide concentration to 25 ppm” (Mineral Resources, 30 CFR, 2014). The purpose of the mine ventilation system is to provide a sufficient quantity of air to the underground mine workings, to dilute methane and other contaminants, to maintain a suitable working environment, and to prevent accidents from happening (Cheng and Wei, 2014). In addition, the organic nature of coal amplifies risks from poisonous gases, explosions, and fires (Engström et al., 2018). If an underground coal mine ventilation system fails after an explosion, the concentration of toxic gases, most notably carbon monoxide, will increase. Carbon monoxide is dangerous because it creates tissue hypoxia by binding to hemoglobin (Özmen and Aksoy, 2015).

Jan 1, 2019

By J. R. Tonry

In the grinding of dry non-metallic materials such as rocks, ores, gypsum, lime, and cement clinker, it is recognized that only a relatively small portion of the energy is utilized for actual size reduction, and the major portion of the energy used is converted into heat. As a result, heat is generated inside grinding mills imparting to the ground product a higher temperature than that of the mill feed. In some cases, as for instance in grinding cement clinker and gypsum into finished cement, it may be desirable to keep this temperature increase within certain limits, to avoid any change in the properties of the product and, because consumers may specify a maximum temperature of the finished product for reasons associated, with the tulimate use of the cement. The actual proportions of heat and useful work are somewhat indefinite. Some tests have indicated that nearly all the energy is converted into heat, and. other investigators have found that they could account for "everything except the noise." Investigations sponsored by the American Institute of Mining and Metallurgical Engineers, and carried out with an adiabatic calorimeter, indicated that a surprisingly high 19 percent of the energy might actually be used for size reduc¬tion. However, for the purpose of the following discussion it will be assumed that: (Fig. 1) 10% is lost in motor and gears 10% is lost by radiation and convection from the mill body 10% is used for size reduction, and. that 70% of the energy input leaves the mill as heat in the ground. product. A typical example (Fig. 2) would be a 7x22-ft tube mill operating in open circuit, driven by a 400 hp motor, charged with 77,000 lb of grinding media and producing 70 bbl or 26,300 lb of cement per hour. Reducing this cement from a clinker size of 96 percent passing a No. 16 mesh sieve to a specific surface of 3,230 sq cm per gm (Blaine) will increase the temperature of the cement approximately 143F above the temperature of the feed, as explained later. If the tempera¬ture of the clinker therefore is 125F, the finished cement will be 268F, which in some cases may be too hot. : When cement temperatures approach a range during grinding where dehydration of the gypsum becomes a problem, a logical consideration would be some means of reducing the temperature before it attains the range conducive to gypsum dehydration. Regardless of the varying degrees of heat generated within the mill or mills during grindingto various finenesses, consideration of the maximum safe limit for any type of cement being ground, which may vary fromplant to plant, govern in refining the mill circuit to obtain a safe temperature range. All plants are not necessarily required to do the same cooling job, and may or may not use cement coolers following the grinding circuit. (Not discussed in this paper). In the past, various means have been employed for controlling this temperature increase, Such means have included water cooling of the outside of the mill shell; equipping the shell with cooling fins; injection of moisture inside the mill; ventilation through the mill; and., in cases of closed-circuit grinding, drawing cooling air through the classifier; the purpose in all such cases being to remove heat from the grinding system.

Jan 1, 1958

By Walter Kahn

Increasing consideration is being given to simplifying traditional methods of tracing pipelines and maintaining required temperatures of stored fluids. Because of their reduction in capital costs, virtual elimination of maintenance. ease of control and "cleanliness"; interest in electrical techniques has accelerated. 1,2 Resistance cable and other heaters have been given some attention. but less well known in the recent introduction of thermal conductive coatings. This paper is addressed to the use of thermal conductive coatings for heating fluid-transport piping and storage tanks. Emphasis will be given to the simplicity, efficiency, and versatility of design and application which their unique properties further contribute to the general economics of electrical heating.

Jan 1, 1965

By W. C. Goldberg

This paper presents pit slope monitoring and back-analysis case studies con-ducted to evaluate and, if possible, stabilize slopes at the Berkeley Pit in Butte, MT. Two specific slopes were studied in detail. The first slope was determined to be near limiting equilibrium, as evidenced by its response to the initiation and cessation of mining, while the second failed on three occasions. Back-analysis of the failed slope verified the laboratory-derived geomechanical properties of the earth materials. Early detection of slope movement can show a need to take action, including modifying slope geometry, dewatering if the slope were wet, and/or investigating and implementing artificial supports. Monitoring slope displacements in open-pit mines can supply data for predicting collapses and preventing serious damage, injuries, and/or loss of life.

Jan 1, 1989

By Tim Neil, O&apos

Large is the applicable adjective here. The largest hydraulic shovel in the world, the largest copper mine in the world, the largest copper producing company in the world. The O&K RH 300 hydraulic excavator has been put into service to move overburden and mine ore at the Chuquicamata open pit mine of the Chilean state copper mining company, Corporation Nacional de Corbre de Chile (Codelco). A large shovel: Equipped with a 26-m3 (34-yd') rock shovel, the RH 300 is able to fill a 170-t (187-st) truck in three passes with cycle times of 30 seconds. A large mine: At present, the Chuquicamata pit is 4 km (2.5 miles) long, 2 km (1.2 miles) wide, and 520 m (1700 ft) deep from rim to bottom bench. The current 450 kt/d (496,000 stpd) of material mined will increase to 600 kt/d (661,000 stpd) when the expanded production rate is reached in 1989. (Chuqui Expansion details, p. 20). A large mining company: Codelco's total output from its two open pit and three underground copper mines is more than 1.1 Mt/a (1.2 million stpy) of copper. This makes Codelco by far the largest copper producer in the world. It accounts for about 15% of the Western World's copper requirements. Chuquicamata is located in the hot and arid Atacama desert of northern Chile. The mine is located 240 km (150 miles) northeast of the port city of Antofagasta. Mine elevation is 2700 m (8850 ft). Chuquicamata's known reserves of oxide and sulfide ores amount to more than 450 Mt (496 million st) of contained copper. There are also reserves of more then 1.4 Mt (1.5 million st) of contained molybdenum. Originally, Chuquicamata was a rail pit. The shovels loaded rail cars hauled by locis on tracks along the benches. During the 1970s, the rail transport was replaced by truck haulage. Crushing and conveying rock waste in the pit is the latest step in the continuing evolution of more productive and economical methods of materials movement. In 1990, the in-pit crushing plant will be in operation. The waste will be mainly transported out of the pit on belt conveyors rather than trucks. The system will incorporate a 9.6-kt/h (10,600-stph) semi-portable crushing station, a 14-km (8.7-mile) conveyor using 1.8 m (6 ft) wide heavy duty belts, and a tripper-spreader to stack the waste dumps. Since Chuqui went into production in 1915, more than 2.5 Gt (2.8 billion st) of material has been excavated. This includes nearly 1.7 Gt (1.9 billion st) of ore from which some 16.4 Mt (18 million st) of copper has been recovered. The Chuquicamata porphyry type ore body is an altered and fractured intrusive granodiorite or quartz-monzonite porphyry that carries the mineralization. The action of ground water on primary sulfide minerals in geo¬logic time has created oxide ores and a zone of secondary enrich¬ment. The oxides occur above the sulfide zone and are also mixed with the upper parts of the latter. At Chuquicamata, the current grade of 1.5% copper is expected to decline to about 1% by 1995. To maintain metal production at cur¬rent levels, it will be necessary to increase ore production rates by about 50%, as mentioned, to 153

Jan 1, 1988

By G. Chi, M. C. Fuerstenau, R. C. Bradt, A. Ghosh

Technical Papers, Mining Engineering Vol. 49, No. 12, pp. 70-75, Discussion by McIvor, R. and Fink, R. P. We must challenge what we consider to be dubious statements in the above-mentioned article. Much of what is reported was published elsewhere over a year ago (i.e., in their reference: Chi, G., Fuerstenau, M.C., Brandt, R.C. and Ghosh, A., 1996, "Improved comminution efficiency through controlled blasting during mining," International Journal of Min¬eral Processing, Vol. 47, p. 93). There is nothing unusual or significant about a 10% or more difference in grindability between two materials of apparently similar origin and other attributes (such as chemical assays, hardness and specific gravity). Consider also that: • the materials came from different locations in the pit; • only two sets of four blastholes were considered per blast: • spatial orientation with respect to the face and other blastholes was also altered; • only two pairs of data were collected; • the impact of sampling error and bias, • both of which could be significant in this case; and • normal experimental error in parameter measurement and calculation. It is clearly inappropriate to draw any conclusions with regard to cause-and-effect relationships without first presenting a thorough error analysis. A substantial number of randomly selected test and control areas might provide meaningful statistical evidence of the claimed effect of the variables being studied. Even that appears doubtful, however, because Figures 3 and 4 show about as much variation in crushing energy for the two "regular" conditions vs. regular and altered conditions. On a separate, but also important issue, broad experience with autogenous/semiautogenous milling has shown that smaller or weaker feed material results in reduced tonnage. ? Reply Thank you for the opportunity to respond to the comments of Messrs. Mclvor ant Fink. The points raised are addressed in the order of presentation. As the commenters noted, there are parallels between this article and a previous manuscript. The reason for these complementary articles is that the audiences for the two publications are different, although there is obviously some overlap. We believe that the different audiences each profit from the technical aspects of the two manuscripts. We are not in agreement that a grindability difference of 10% is insignificant. The laboratory work is corroborated by data obtained at plant scale, which confirms that an increase in throughput of the same magnitude (between 5% and 10%) is achieved when the blasting pattern is modified, i.e., reduced from 9 to 8 m (29.5 to 26.2 ft). This increase in throughput translates into substantial power savings for ore grinding. No differences in appearance in the rock were noted between the two adjacent areas. Rock in both areas belonged to the same rock mass/block, which was quite homogeneous and without geological features in between that could create drastic differences in comminution characteristics. Given the resources, we would have sampled from all of the holes. Additional data would certainly have been desirable, as would have been additional funding to expand the study. Their comment, "spatial orientation with respect to the face and other blastholes was also altered," is not clear. Figure 1 does not show the initiation sequence. Without the initiation sequence, we do not understand how the commenters concluded that spatial orientation with respect to the face and other blastholes was altered. The commenters may be trying to make the point that the four-hole configurations with respect to the original bench face are not exactly the same. In fragmenting the

Jan 1, 1999

By K. A. Smith, X. Cheng

Introduction The separation of minerals from complex sulfide ores by flotation requires careful control of the surface chemical behaviors of the minerals constituting the ores. When the minerals in the ore are finely disseminated, extremely fine grinding is needed to liberate the minerals. Fine grinding is not only an energy-intensive operation, but also influences the flotation behaviors of liberated minerals. Steel balls are commonly used for grinding. Galvanic interactions occur not only between sulfide minerals and grinding media but also among the different sulfide minerals themselves. The electrochemical characteristics of chalcopyrite and pyrrhotite when they are independently coupled with steel grinding media and the relevance of these characteristics to the flotation behavior of these minerals have been exten¬sively studied (Cheng and Iwasaki, 1992: Li and Iwasaki, 1992). These studies dealt with reactions in two electrode combinations, simulating mineral-grinding medium and mineral-mineral interactions. In practice, however, it is more common to involve more than one sulfide mineral. In a study of a pyrite-pyrrhotite-mild steel system, Pozzo et al. (1989) provided evidence that a mineral having an intermediate potential could behave differently, depending on the condi¬tion of galvanic contact. This investigation was undertaken to study a chalcopyrite-pyrrhotite-mild steel system to gain better in¬sight into the mechanism of multiple galvanic interactions. A three-electrode system was chosen to simulate a typical copper sulfide ore. The sulfide minerals were represented by chalcopyrite and pyrrhotite. The grinding medium was rep¬resented by mild steel. The nature of galvanic interactions involving these minerals and grinding medium in an oxygen atmosphere was investigated in a three-electrode combina¬tion, and an attempt was made to relate the laboratory flotation results of pyrrhotite-chalcopyrite ores. Possible implications to the differential flotation of complex sulfide ores are discussed. Key words: Electrochemistry, Flotation, Sulfide ores Experimental The pyrrhotite (57.74% Fe, 0.09% Cu, 4.84% Ni and 36.54% S) sample was obtained from Falconbridge Ltd.'s Strathcona mine, Ontario, Canada, and the chalcopyrite (26.00% Cu, 28.57% Fe, 0.05% Ni and 31.52% S) sample came from Transvaal, South Africa. Mild steel (AISI 1020, 0.2% C) was chosen to represent grinding media. A conventional three-electrode cell was used in the elec¬trochemical measurements. The working, counter and refer¬ence electrodes were, respectively, sulfide minerals or mild steel, graphite rod and saturated calomel electrode (SCE, 0.245 V vs. standard hydrogen electrode). The exposed areas of the sulfide minerals and grinding medium electrodes were approximately the same (0.17 to 0.19 cm2). The preparation methods of working electrodes, the reagents used, the experi¬mental procedures and the instrumentation were the same as those described previously (Cheng and Iwasaki, 1992). Rest potentials of pyrrhotite, chalcopyrite and mild steel were measured individually in 0.05-M sodium sulfate solu¬tions. Combination potentials and galvanic currents of the sulfide minerals and grinding medium were measured in short-circuited two- and three-electrode systems. All poten¬tials in this study are quoted in volts with respect to the standard hydrogen electrode (SHE). To simulate the constantly abrading conditions in a ball mill, rest potentials, combination potentials and galvanic currents were also measured by rotating a porcelain ball against chalcopyrite, pyrrhotite and mild steel electrodes in

Jan 1, 2000

By D. A. Norrgran, J. A. Wernham

Introduction In view of the continuing depletion of natural resources, requirements for energy conservation, plant optimization and hazardous waste management, and environmental consciousness, recycling and secondary recovery of materials are being developed with increasing interest Many secondary recovery processes are being developed that apply mineral processing technology. Gravity separation, magnetic separation and flotation have all emerged as explicit methods for recovering residual values from various process streams and hazardous constituents from waste streams. Recently, a technique for concentrating metallics has been successfully reintroduced. The concept of the alternating magnetic field eddy-current separator was developed over a century ago using rudimentary electromagnets (Patent 400317 issued to Thomas Edison on March 26, 1889). Widespread acceptance was, however, hindered by high costs, marginal results and poor economic incentives for secondary recovery. With the recent increased interest in secondary recovery and the development of high-energy rare earth permanent magnets, the eddy-current separator has now gained prominent acceptance. The separation forces imparted by an eddy current separator are the function of the magnetic field intensity and the alternating frequency of the magnetic field. The evolution of permanent magnets has provided a cost effective alternative for the generation of high-intensity magnetic fields. High-energy rare earth permanent magnets have substantially reduced the capital and operating costs over the electromagnetic circuits employed in antecedent eddy-current separators. Specifically, in recent years, the strength of permanent magnets has increased several fold with neodymium-boron-iron rare earth magnets now providing an energy product of 35 million gauss-oersted. Figure 1 shows the evolution in the strength of permanent magnets. The development of these rare earth magnets has led to the design of circuits possessing a magnetic force an order of magnitude greater than that of conventional permanent magnetic circuits. The eddy-current separator has been successfully applied in several metal sorting and recovery operations. Most common is the sorting of metal from shredded automobile scrap and municipal waste. With the advent of high-frequency, high-intensity magnetic fields, the separator has advanced to the point where it has direct application in the beneficiation of fine sized metals. Relatively fine sized (-1/4 in.) metal bearing slags and spent foundry casting sands as well as precious metal hearing electronic scrap demonstrate excellent metallurgical response to the eddy-current separator, and applications have progressed to finer materials and more selective separations. Theory Electrical currents are induced in all conductors when exposed to an alternating magnetic field. The induced current generates a magnetic field in the conductor that opposes that of the alternating magnetic field. Figure 2, is a schematic illustration of generated eddy currents. In this particular example, the alternating magnetic field is produced by a series of permanent magnets mounted on the circumference of a rotor. The permanent magnets have alternating polarity. As the rotor revolves, an alternating magnetic field is produced, and the rate of revolution determines the frequency of the alternating magnetic field. When a conductor, such as a metallic disc in the illustration, is placed in the alternating magnetic field, a closed loop current flow occurs. The current loop in the conductor produces a magnetic field that is a mirror image of the alternating magnetic field. At any point in time, the induced magnetic field in the conductor directly opposes the alternating magnetic field. This opposition of magnetic fields produces an instantaneous repulsion in the conductor. The conductor is consequently repelled from the rotor. The alternating magnetic field has no effect on nonmetallic materials as they pass through the magnetic field and discharge the rotor in a natural trajectory. [ ]

Jan 1, 1992

By Don S. McCoy, W. Joseoph Schlitt, Donald B. Swanson

Introduction Contract miners have long played an important role in the minerals industry. Historically, these were individuals who worked as underground miners and were paid by the ton or the day. These hardy types still ply their trade in the industry. However, today the term "contract miner" generally refers to an independent contractor who supplies the people and equipment needed to conduct surface mining operations for an owner or lease holder. At one time or another, contract miners have probably handled virtually every mineral commodity from sand, gravel, and stone quarrying to mining uranium and precious metal ores. In today's market, though, contract mining appears to be concentrated in just a few areas. These include precious metal mines, various types of solid fuel operations, and reclamation of abandoned mines. This article covers market areas, contract mining organizations and services, and their advantages and disadvantages. Gold and silver mines Although exact figures are not available, contract mining represents a significant element in the precious metals segment of US mining. For gold and silver, the growth in contract mining appears to be paralleling the overall growth in mining activity itself. Nevada is the largest gold producer in the US. The latest available data (1985) on open pit gold/ silver mines is summarized in Table 1 (Nevada Division of Mine Inspection, 1986). These figures show that contractors mined nearly half of the ore in 1985. Total tonnage mined, of course, was much higher for both owners and contractors. Assuming an average stripping ratio of 2.5 to 1, the total daily mining rate for the contractors was more than 145 kt (160,000 st). This translates into a contract mining rate of about 36 Mt/a (40 million stpy) based on 250 operating days a year. This approximate 50-50 split between owner and contractor also appears to hold for the future. Metals Economics Group (1987) published a survey of 51 short-range and 54 long-range precious metal projects being considered throughout the US. Contacts with owners, requests for proposals, and past corporate philosophies all suggest that about 55% of the open pit operations scheduled for the near-term will be contract mined. This represents an estimated 26 kt/d (28.5 stpd) of ore. A similar assessment of the long-term projects indicates that about 12 kt/d (13,000 stpd), or just under 50% of the open pit tonnage will be contract mined. There are several reasons why gold and silver represent such an attractive opportunity for contract miners. One is that many properties are owned or leased by individuals or partnerships that are essentially entrepreneurial. Many of these owners or lease holders have minimal mining experience. They must therefore either rely on contract mining organizations, sell an interest in the project to another mining company that will become the operator, or hire an experienced management team. For about the same reason, junior mining companies also represent good prospects for contracting. This is because most junior companies focus on exploration. They would prefer to spend their generally limited funds on drilling programs. Thus, they tend to minimize additional capital requirements by using contract miners or farming out the property to another mining company. Another good reason for using contract miners is the size of the deposit. Many properties are so small that their reserves will not support the capital expenditures

Jan 12, 1987

By H. G. Naidu, Raj K. Singhai, S. P. Singh

Introduction The economics of open-pit mining de¬pend heavily on the cost of overburden removal. It is necessary to maintain a pit slope as steep as possible. Steeper slopes lower the volume of overburden to be handled, thus lower the cost of material handling. There is, however, a limit to which the pit slopes in any given situation can be steepened. Steepening beyond this point could jeopardize the pitwall stability, ren¬dering it unsafe. It might even result in curtailment of pit production. Several factors affect the stability of pitwalls. Some are natural, such as topogra¬phy and regional and local geology. Others are man made, such as pit geometry, ground water management, equipment, drilling and blasting, and the type of excavation. Greater control can be exercised in the latter with proper planning and by following a set of procedures based on field experiences. This article deals with one such control¬lable factor - drilling and blasting. The eco¬nomics of steeper pit slopes is discussed, followed by a record of drilling and blasting practices in western Canadian surface mines. Microprocessor-based instrumentation and its application to blasthole drilling is also described. Economics of steeper pit slopes The economics of steeper pit slopes is not a new issue. It should be reinforced, though, that pit slopes must be as steep as possible to ensure maximum recovery of minerals at a minimum stripping ratio. For operating, economic, and safety reasons, different pit slopes are maintained during different phases of pit development. Figure 1 shows a hypothetical ore body, cylindrical shape, with a radius of 225 m (740 ft) and located 30 m (100 ft) beneath the surface. The mining operations are planned at two different slope angles of 25° and 40°. The ultimate pit depth was established on the basis of the criterion of "break even pit." The variation of the stripping ratio at differ¬ent mine levels is shown in Fig. 2. Table I summarizes the eco¬nomics involved for two pit slopes, 25° and 40°. Figure 3 presents the profits during the life of the mine for these two pit slopes. This study shows that steeper slopes for an open pit operation are far more advanta¬geous than flatter slopes. But, what is the optimum slope at any one point in the life of a moving pit? And how can the stable slopes be maintained over the short- and long-term? Canadian blasting practices A survey of blasting practices in western Canadian surface coal mines revealed that, at most mines, hard consolidated overburden is drilled and blasted. Blastholes are drilled with rotary crawler-¬mounted rigs using bits in the 220 to 310 mm-diam (8.5 to 12 in.-diam) range. Typi¬cal hole depths are 12 to 18 m (40 to 60 ft) to accommodate the pit bench heights of 12 to 15 m (40 to 50 ft). For a 12-m (40-ft) high bench, a typical hole will be 14.6 m (48 ft) deep including the subgrade drilling of 2.6 m (8.5 ft). Half of this hole will typically be a powder column and half will be stemmed. For a 15-m (50-ft) bench, a typical hole will be up to 18 m (66 ft) deep including the subgrade drilling. Drilling patterns are usually 9 m (29.5 ft) spacing and 9 m (29.5 ft) burden. De¬pending on local geology and hole diameter selected, drilling patterns of 8 x 8 m, 8.5 x 7.3 m, and staggered patterns of 11 x 5.5 m

Jan 1, 1989

By B. Beatty

Discussion by B. Beatty This is a response to M.J. Hrebar's paper, which tracks in detail the highly productive stripping capabilities of modern large bulldozers at two separate installations. It proves the old adage "economies of scale" and quantifies the advantage of being able to move large loads of dirt over comparatively short haul distances by sliding rather than carrying the load. This discussion examines whether a bulldozer is the most efficient way of using this geotechnical principle. Case 1 - Study eastern US: Contour coal stripping application The energy consumption for the Case 1 mining method (eastern US contour stripping operation) shows the D11 averaged 117 L (31 gal) of diesel fuel per scheduled hour, and produced 333.3 bank m3 (436 bank cu yd) per scheduled hour (0.351 L/m3). Assuming 1 L (0.26 gal) of diesel fuel contains 10.4 kW-hr of energy, the unit expenditure of fuel energy for Case 1 Study was 3.66 kWh/m3. Another way of shifting this material is by using an overburden slusher (OS), as shown in Fig. 1. The OS as shown in Fig. I. The OS is electrically powered and consists of a manned low wall winch (LW ), a remote controlled (radio telemetry) highwall winch (HW) and a hoe. The winches operate to pull the hoe back and forth across the pit. The hoe carries load in the forward direction ahead of a rear mounted blade and returns empty. A spreadsheet (Lotus 123) simulation package, BOSSIM, models the OS operation on a load by load basis over a wide range of pit configurations. The geometry shown in Fig. 2 of the paper was modeled using BOSSI M and the results are summarized in Table1. Table 1 shows a LW with 150 kW (200 hp) and a HW of 120 kW (160 hp) installed power, towing a 15-m3 (19-cu yd) hoe, would accomplish the Case 1 stripping task at about the same scheduled stripping rate (D11=333 m3/hr or 433 cu yd/ hour, OS=342 m3/hr or 447 cu yd/ hour). It was assumed the winches in this equipment have an electromechanical efficiency of 90%. This gives an average electrical energy expenditure rate of 151 kW-hr, and a unit energy expenditure level of 0.44 kWhr/m-3(0.57 cu yd/kWh)-an improvement of 8291%- over the D11 alternative. The energy comparison illustrates the importance of using more efficient electric power if possible, ahead of the internal combustion engine, and delivering power from a fixed ground position rather than through a moving track system. Another important energy observation is the comparative action of the dozer blade and the OS hoe. The former implement continuously digs and spills as it works the load forward. The OS hoe digs the load until full, the blade then lifts due to the contained load pressure and a shear plane forms under the mass, enabling it to slide forward in a force reduced - transporting mode. Economic comparison Table 2 operating cost per scheduled hour and unit cost comparison follows a format used in Tables 2 and 3 of Hrebar's paper and compares Case 1 Study with the OS alternative. Drill and blast costs are kept constant in both cases, although when larger OS units are considered, it is expected these costs will reduce because bigger hoe sizes will permit larger rocks to be transported efficiently.

Jan 1, 1995

By V. F. Hollister

Although it is apparent that porphyry copper deposits may have formed in the Andean orogen from Permian time onward, commercial deposits thus far developed appear to have Cenozoic ages only. Following Chilean useage, the porphyry deposits are divided into either: (a) a stockwork, or (b) a tourmaline breccia model. Stockwork type deposits tend to occur on major strike-slip structures which may be either transcurrent or transverse faults related to plate movements, while the tourmaline breccia type may more frequently form independently of regional faults. Stockwork types may be conjugate fractures that intersect and ramify as a network of veins and veinlets, cutting altered rock, and genetically related to a major regional fault, or in other cases they may reflect the intersection of two large fault structures. Circular fracture patterns may more rarely occur in stockwork deposits in response to a specific intrusive. Tourmaline breccia types are characterized by circular and radial fractures genetically connected in some way with tourmaline breccia pipe development or an intrusive piercement that preceded it. Each structural model contains mineralized intrusive8 which have a close time-space relationship to ore and in each model, mineral paragenesis is similar, regardless of age or location in the plate. Also, each model contains the same type of alterntion halos (phyllic, argillic, prophllitic) as are developed in intru- sive centers in porphyry copper deposits found in North America. The potassic zone is exposed best in the older deposits. Examples of each type of deposit are illustrated, demonstrating the feature evolved.

Jan 1, 1973