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By C. P. Lazzara, F. J. Perzak, C. C. Hwang, C. D. Litton

Fire development and spread along conveyor belts in ventilated ducts were investigated experimentally and theoretically. Various types of conveyor belts used in mining applications were ignited and burned in a full-scale gallery under various flow conditions. The flame spread rates depended strongly on the ventilation air speed. The time-temperature traces of gas and belt obtained at 12 feet from the entrance show that a rapid belt surface temperature change takes place prior to a change in the gas temperature, indicating that radiative heat transfer from the flame to the belt surface plays an important role. Further, anomalies in the flame spread rate in the vicinity of 1.5 m/s air speed (i.e., much greater rate than that observed at either higher or lower speeds) can be explained through the mechanism of flame tilt and the resultant view factor between the flame surface and the belt surface. In order to correlate the fire spread with material properties of the conveyor belts (such as the specific heat, the thermal conductivity, and the pyrolysis temperature) and the fire environment (such as the gas temperature and the ventilation air speed), a theoretical model has been developed. This model is based on the experimental observation that the radiative heat transfer is the dominant mode of energy transfer ahead of the flame front. Agreement between the theory and the experimental results is good.

Jan 1, 1991

By N. Colvin, J. W. Gulyas

HISTORY The Surigao Nickel Refinery of Marinduque Mining and Industrial Corporation is located on Nonoc Island off the northwestern coast of Mindanao in the Philippines. Nonoc Island is part of the Surigao Mineral Reservation which had been identified as early as 1912. Initial interest in these deposits was due to their iron content, and their value as a nickel deposit was not seriously entertained until the late 1950's. The Philippine government undertook extensive test pitting and sampling programmes at that time and in the early sixties Marinduque became involved in helping to establish firmer ore reserves. Evaluations of the test pitting programme indicated reserves of 75,000,000 tonnes of ore averaging 1.22% Ni, 0.1% Co and 38.5% Fe. The Philippine government called for bids on the reservation in 1962 and Marinduque and Sherritt Gordon carried out some feasibility studies for the recovery of nickel from the deposit. However, it was not until 1968 that Marinduque was granted proper authority to exploit the Surigao deposits. Marinduque and Sherritt then entered into an agreement which allowed Marinduque to utilize a process developed by Sherritt for treating lateritic ores. During 1968, a 25 st/d pilot plant was erected in Fort Saskatchewan, Alberta for the purpose of modelling all process steps. A 10,000 tonne ore sample was sent from Nonoc as feed stock for the pilot plant.

Jan 1, 1979

By B. J. I. Adair, K. R. Weller

The bulk scavenger tailing from Aberfoyle's Hellyer Concentrator, which represents about one-third of the total tailing, runs about 2.5-3 g/t gold and 140-200 g/f silver. Both values are very finely disseminated, particularly the gold. Aberfoyle wished to establish whether ultra-fine grinding followed directly by cyanide leaching could be applied to the tailing for recovery of gold and silver. The work comprised the determination of gold and silver recoveries as a function of grind size and leaching time, the specific energy consumption required for continuous wet grinding of the tailing to the specified product size in a small horizontal continuous stirred mill, and an estimate of the rate of media consumption. The maximum extraction of gold, 23.3% was obtained from the finest product, with the highest cyanide concentration and longest leach time, confirming earlier work. A similar result was obtained for silver. The specific energy requirement for milling rose from 27.7 kWh/t for a P80 of 8.1 um to 219 kWh/t for a Pea of 2.3 um. The consumption of ceramic media was about 1.6 g/kg of final product. An extensive QEM'SEM Trace Mineral Search on the unground tailing failed to identify any gold-bearing particles but did identify silver-bearing particles which displayed a variety of textural associations and liberation characteristics.

Jan 1, 1998

By John A. DesLauriers, Hans J. Broennimann

INTRODUCTION The early record of tunnel excavating machines in hard rock tunneling has not been one to particularly brag about. We have all heard the stories of major difficulties when trying to break hard rock with machines at promised advance rates of several times that of conventional drill and blast. Unfortunately, sometimes this has not happened. It has been through this trial of the past generations disappointments that today's TBM's have evolved to their present high level of performance in a wide range of ground conditions. Today's machines offer much increased reliability and durability. This translates into high availability to the tunnel contractor. Although there still are occasional disappointments being reported, the state of the art is that reliable and predictable machine excavation of rock now is both a reality and well founded expectation. The most recent editions of Atlas Copco's MINI/MIDI Fullfacer, we feel, make a strong contribution to the reliable performance of todays TBM's---especially in small-size rock tunnels. This report will, 1.) explain the equipment and effect of rock both on the cutting technique and machine performance; 2.) list advantages of the non-circular shape, during and after construction; 3.) list applications in civil and mining industries; 4.) discuss North American experiences, performance and costs; 5.) list advantages over drill and blast and; 6.) discuss briefly lease arrangements.

Jan 1, 1979

By Nick Barton, Harald Hansteen

The deformations resulting from excavation of very large openings are compared using two-dimensional F.E. continuum analyses and discontinuous physical models (20,000 discrete blocks). Both the joint orientations and the model horizontal stress levels were varied. Some models were dynamically loaded to simulate earthquakes (0.2-0.7 g). Model deformations were recorded using photogrammetry. The changes in deformation when increasing the simulated spans from 20 m to 50 m were of particular interest. High horizontal stress caused surface heave when joint orientations were favourable for arch stability. Joint orientations also determined whether the pillars between parallel openings were in a state of compression or tension. INTRODUCTION The engineering performance of large rock caverns has traditionally been learned from mining and hydro power projects, where the depth below surface is often many times greater than the span of the openings. Deformations measured in the walls and roofs of hydro power caverns generally range from about 5-50 mm, though there is a documented case where a wall moved in 126 mm (1), and another where the arch moved down 147 mm (2). The chief objectives of the present studies of large near-surface openings were threefold: (i) to provide deformation data to compare with monitored data from planned engineering projects involving large span near-surface excavations, e.g. underground sports complexes, civil defense shelters, nuclear power stations,

Jan 1, 1979

By E. Ozberk, J. B. Hiskey, M. J. Collins, B. J. Scheiner, M. E. Chalkley, I. M. Masters

Conventional zinc concentrates generally contain 5 to 10% Fe, while bulk concentrates can contain up to 20% Fe. As a result, the disposal of iron bearing residues is an integral part of zinc refining operations. The quality of iron residues must comply with ever changing and increasingly stringent environmental legislation. Zinc has been recovered from sulphide concentrates by a hydrometallurgical route for almost one hundred years. The disposal of iron was a major difficulty for the industry for many decades and was directly responsible for low overall zinc recoveries. The introduction of the jarosite process in the mid 1960's allowed for the treatment of neutral leach residues and the precipitation of iron in an easily filterable form. Subsequently, three additional iron precipitation processes have found commercial application in the zinc industry: the goethite process; the paragoethite process; and the hematite process. The Sherritt Zinc Pressure Leach Process has been developed for integration with existing roast leach electrowin zinc plants, or as a stand alone facility for the treatment of a wide variety of zinc bearing feed materials. It is essential that these processes allow for the disposal of iron residues in an environmentally acceptable form. This paper describes the integration of the Sherritt Zinc Pressure Leach Process with commercially proven iron precipitation processes. The results of laboratory testwork and data from operating plants are presented.

Jan 1, 1993

By M. Gobla

"Since 1994 the Bureau of Reclamation has used risk analysis to evaluate dams. Risk considers both the likelihood of failure and the expected consequences should failure occur. Risk analysis relies on the detailed formulation and critical analysis of all possible potential failure modes. Potential failure modes are formulated by describing the sequence of events that must occur to result in failure, where failure is defined as the uncontrolled release of the impounded reservoir. Failure occurs in response to loadings such as water seepage through a dam (static forces), shaking from an earthquake (dynamic forces), sudden flood inflows (hydrologic forces), or accidental overfilling (operational errors). Reclamation has applied risk analysis methods to inactive and abandoned mines to evaluate tailings dams and flooded underground mines. Case histories of potential failure mode analysis as applied to mine impounded water illustrates what critical technical information is needed to determine if failure is likely, what the consequences of failure are likely to be, and what are the highest risks that should be addressed first. INTRODUCTION Impounded water, and high water content wastes, at mining and industrial processing sites are potential threats to public safety and the environment. Recent failures such as the TVA Kingston flyash impoundment failure in Tennessee in 2008 (5.4 million yd3 released), the Mt Poly Tailings Dam failure in British Columbia in 2014 (5.8 million yd3 released), and the Smarco Fundao Tailings Dam failure in Brazil in 2015 (78 million yd3 released) has renewed attention to this issue. Potential risks from impounded water at mining and industrial sites originates not only from tailing dams, but also from water in process ponds and conveyance pipelines, sediment retention dams, water retention dams, mine pit lakes, and flooded underground mine workings. Conventional water retaining dams also have a record of failure. Notable examples of failure include Vajont Dam in Italy in 1963 (2,600 lives lost), Banqiao and Shimantan Dams in China in 1975 (171,000 lives lost), and Machhu-II Dam in India (10,000 lives lost). After the 1976 Teton Dam failure in Idaho, the Bureau of Reclamation (BOR) established a Dam Safety Program. By 1984 the agency developed and began using risk analysis methods. Reclamation operates and maintains an inventory of hundreds of aging dams which were designed and constructed over a period of one hundred years. Many of these facilities were constructed using engineering design and construction methods which would not be considered good practice according to present day standards. Diligent observational monitoring, detailed site investigations, and the application of risk evaluation to the BOR dam inventory has identified many deficiencies resulting in significant demands on available funding for remediation efforts. The application of risk analysis has allowed the agency to rank sites by risk. By giving priority of remedial efforts to high risk sites, significant improvements to dam safety have been achieved. BOR has also applied risk analysis to evaluate and reclaim inactive and abandoned mining and industrial sites. The application of risk analysis methods in the mining industry is not new, the methodology was transferred to the mining industry more than 15 years ago as discussed in ICOLD Bulleting 121 (ICOLD, 2001), yet failures persist. This paper is meant to convey the lessons learned from BOR’s use of risk analyses in its work on mine sites."

Jan 1, 2017

By M. Szatkowski

The studies conducted by Laplante and coworkers (1983,1984) with flotation of galena and quartz, showed that the rate of flotation reaches a maximum value for a certain intermediate value of the air flow rate. A similar, though less pronounced, effect of the air flow rate on the selectivity was also observed. Luttrell and coworkers (1985) showed that the size of air bubbles has a significant effect on both rate and selectivity of coal flotation. These two studies used flotation tests in which no froth layer was formed. This note describes effects of the air flow rate and bubble size on the rate constant and selectivity of coal flotation, using a mechanical laboratory machine with a froth layer allowed to form. Experimental A raw, dry ground Illinois coal was used in flotation tests. The size distribution and ash content of the samples are : Particle size (µm) Weight (%) Ash (%) +600 4.30 41.7 600/200 43.05 19.0 200/74 22.25 16.0 - 74 30.40 24.0 average 20.8 No. 2 fuel oil with a concentration of 200 g/t of feed and M.IBC with concentrations of 11, 22, and 33 mg/L of pulp (128, 256, 384 g/t, respectively) were used as flotation reagents. Tap water with a natural pH of 7.0 to 7.8 was used throughout the tests. No other reagents were used. The starting concentration of coal in the pulp was 86 g/L. A standard flotation machine with the pulp volume of 5 L was used in experiments. Three different air flow rates were employed, resulting in three air escape velocity values of 14, 30, and 70 cm/min. The air escape velocity (AEV) was calculated as the ratio of air flow rate (Qa) to the slurry surface (Ss) in the machine. For a comparison, the upper limit of air escape velocity values used in commercial size flotation machines is between 70 and 100 cm/min (WEMCO, 1983). A machine consisting of a generator of fine bubbles and a separatory column, described elsewhere (Szatkowski and Freyberger, 1985), was used for some tests where bubble size lower than that generated in the standard machines was required. Prior to a flotation test, a batch of coal was mixed with water in the flotation cell for at least 10 minutes to ensure that all coal was thoroughly wet. Then the specified amount of fuel oil was added and the pulp was conditioned for five minutes. The specified amount of MIBC was added 15 seconds before the end of conditioning. After air flow was opened, flotation products were collected for 0 to 30, 30 to 60, and 60 to 120 second intervals. Content of the combustible material in samples was determined by subtracting the ash content from 100%. A photographic technique was used to determine initial distributions of bubble size generated by the flotation machines. Bubbles generated in water containing the prescribed amount of the frother but no solids were photographed against a black background. The average diameter and standard deviation were then calculated for each set of flotation conditions. The method gives results reproducible within k 5% of the average bubble size. Table 1 contains information about the measured bubble size distributions expressed by the average bubble diameter (Db) and the standard deviation ( SD ) . Results and discussion The general formula for calculating the flotation selectivity is: o = recovery of coal/recovery of ash (1)

Jan 1, 1988

By C. J. Veal, C. I. House, S. R. Bellamy

INTRODUCTION Coal gold agglomeration (CGA) is a new concept in gold processing. It involves the recovery of oleophilic (oil-loving) gold grains by causing them to adhere selectively to the surfaces of coal oil agglomerates. Under optimum conditions very high recoveries of gold. can be achieved. The agglomerates are recycled to build up gold loading to typically 1000 - 5000 g/tonne. Bullion can be produced by burning the loaded agglomerates with subsequent recovery of gold from the highly loaded ash either by dissolution/electrowinning or smelting. A wide-ranging programme of laboratory testwork has been performed to investigate the response of different ore types to the process. This has shown CGA to be most applicable to ores in which the gold is liberated or easily liberatable, relatively coarse (ca 500 µm) and fine (<5 µm) gold grains being recoverable with equal ease. The best opportunities for CGA seem to be: • placers containing fine gold not recoverable by gravity concentration • low grade gravity concentrates to avoid multiple magnetic/gravity separation stages and amalgamation • free milling ores where the gold can be liberated at a reasonable ore particle size eg >d80 50 µm • tailings past or present from gravity concentration plants. CGA testwork ranging in scale from laboratory to pilot plant on tailings produced by gravity concentration in the 19th century Australian gold rush has already been reported (Bonney, C F, 1988; House, C I, Townsend, I C and Veal, C J, 1988). An economic appraisal for a Canadian free milling gold ore has also been published (Buckley, S A, House, C I and Townsend, I G, 1989). This paper describes results of a detailed laboratory programme to assess CGA as a means of recovering gold from a beach sand deposit in New Zealand. The area from which the samples were taken consists of quarternary interglacial marine, lagoonal, fluvial and eolian deposits. The heavy minerals present are predominantly ilmenite and garnet and small concentrations of gold. In certain locations the sands and gravels are consolidated by a limonitic cement. In the 19th Century some parts of the deposit were mined and the ore processed using stamp milling and gravity concentration. Gold recoveries were low, probably below 50%, due primarily to the fine size of the gold grains, the flaky nature of the gold and also imperfect liberation by the stamp mills. Our laboratory testwork was divided into three parts: i) Preliminary studies on an initial sample to assess the feasibility of treating the ore by CGA ii) Extended studies on further samples from various locations in the deposit to confirm predictions from (i) iii) Conventional gold recovery techniques. EXPERIMENTAL PROCEDURES All CGA tests were performed by agitating a slurry of ore in water in a 0.6 dm3 pyrex beaker fitted with four vertical baffles. When required, potassium amyl xanthate collector was added and the slurry conditioned for 5 min. Coal (ex BP Coal Australia&apos;s Tahmoor Colliery) previously ground to a d80 50 µm was added followed by a commercial grade of gas oil (density 840 kg/m3). The agglomeration time was 30 min. The agglomerates were separated from the tailings by screening, both were then dried and analysed for gold by fire assay. This test procedure provides a reasonably accurate estimate of gold recoveries that might be expected from our 1 tph pilot plant. Ore samples were ground either dry in an agate ring disc mill or wet in a steel batch tumbling

Jan 1, 1989

By Weixiang Meng, Zhao Song, Ming Guo, Zhigao Liu, Shoubao Zhang

"There are many multi-layer coal reserves in China which account for nearly 20 to 25% of the existing proven reserves. When the coals are mined in a descending order using the fully mechanized long wall mining method, they usually have roof collapse with cavity and air leak in roadways and long wall stopes; this is especially relevant when the thickness of the rock between two coals is less than 5.0 m, as the roof stability of lower seam would be greatly reduced. At the Wuhushan Mine, such abnormal strata behavior occurs, including abnormal support stress, irregular period roof pressure, and roof falls in the face. The main reasons for these behaviors include smaller roof thickness, lower rock strength, and mining influence of upper coal; these reasons were determined by analyzing the in situ conditions. A mechanics-based model of the mine geometry was set up to study the sources of stress, the stress evolution of coal, and the discriminant method of roof stability; to obtain the smallest safe thickness of roof between two coal layers, numerical modeling was applied to analyze the relations between the roof thickness and its displacement distribution. Based on the above research, the No.1001 face was divided into easy roof caving zone (ERCZ), roof caving transition zone (RCTZ), and unlikely to roof caving zone (URCZ). In the ERCZ, comprehensive grouting to reinforce the roof was implemented; in the RCTZ, local grouting to reinforce the roof was implemented. Furthermore, a face stability control index and its range were developed in a fully mechanized face combined with mine pressure observation. After implementing this strategy, there have been no unexpected roof caving or other serious strata behaviors.INTRODUCTIONChina is the country with the largest coal production in the world, and the conditions for its coal seams are particularly complicated. For example, in China’s Datong, Wuda, Pingdingshan, and Qinshui mining areas, there are a large number of multi-layer coal seams. The thickness of these coal seams ranges from 1.3 m to more than 10 m, and the spacing between layers is from under 1.0 m to over 30 m. A common problem with these coal seams is that one layer of mining has a large impact on other coal seams, which are customarily referred to as close-distance coal seams. According to statistics, close-distance coal seams account for about 20–25% and 15–18% of China’s coal reserves and production, respectively.From the reference research of Huang (1999, 2002, 2018) and Tu (2010, 2011) we can learn that in the mining process of close-distance coal seams, downward mining is mainly used. When the upper-layer coal is mined, it has a huge mining disturbance effect on the floor (that is, the roof of the lower coal). The mining disturbance increases the mining fissure in the rockmass and weakens the strata strength. The mining problem formed by this issue is that when the lower coal is mined, problems such as difficulty in roadway support, collapse of stope roof, and difficulty in ventilation are often encountered. Especially for the mining of medium-thick coal seam groups with a layer spacing of less than 5.0 m, there has been a large-scale roof collapse accident with a fall height of more than 10.0 m. In response to such problems, domestic and foreign scholars and engineers such as Tan, Zhao, and Xiao (2010), Xie et al. (2017), and others have also carried out targeted theoretical and practical research."

Jan 1, 2019

By Herschel H. Potter

While preparing this paper, I soon realized I had a problem in defining what injuries should be classified as "personal." Admittedly, in some cases, falls of persons, sprains and strains, pulling and pushing injuries, and back injuries may have been caused by the individual. But, I believe the number of injuries directly attributed to an individual&apos;s carelessness is relatively small in comparison to the total number of injuries in those categories. For discussion purposes, the injuries most often considered to be personal, that is, caused by individuals are those classified as back injuries. My paper will be devoted to back injuries because this category represents the greatest number of injuries occurring in coal mining. A man&apos;s best friends are his wife, his back, and his dog. The back, however, has a reputation of not being as faithful as his other two friends.1 "Industrial back injury is one of the major occupational health problems plaguing industry to- day. According to national surveys, back injury accounts for more lost time than any other occupational injury and costs companies thousands of dollars in compensation, medical fees and over- time pay."2 A study conducted by USA Today found that back ailments afflict an estimated 75 million Americans, have totally disabled about 2.5 million adults, lead to as many as 200,000 required surgeries each year, are a leading cause of absenteeism and disability, are the second leading cause of hospitalization, are responsible for 18 million doctor visits a year and cost about 93 million work-days annually. The cost to industry is estimated at $14 billion a year.3 During the past 30 years, a great deal of accident prevention education has been directed toward the reduction of back injuries. However, this effort has not produced any great change in the incidence of the injury. Back injuries should be of the utmost concern to everyone here today. I don&apos;t need to tell you what the incidence rate is for back injuries, nor will I use any statistics to inform you of the incidence rate for the coal mine industry or any specific mine. It is only important to know that the number of back injuries is excessive. Back in- juries resulting from lifting have been and continue to be one of the most constant and undiminishing problems in the occupational environment. The fact is, we use our backs to per- form work that the back was not designed to do. Historically, manual materials handling related accidents have consistently accounted for the largest percentage of all industrial accidents. In fact, over 25 percent of all industrial accidents are related to handling of materials.4 In underground coal mining operations, material handling accidents account for even larger percentage of lost time injuries. In 1983, for example, they accounted for 34 percent of all lost time injuries and 32.6 percent of all non-lost time injuries. 5 I want to use one example of a task that will illustrate what I have just said: the building of a concrete block stopping. Each solid concrete block will weigh from 60 to 70 pounds. To build a stop- ping 6 feet high and 20 feet wide will take approximately 80 blocks. On the average, these 80 blocks will be handled at least four times from loading into the conveyance for transportation to being placed while constructing the stopping. This generally requires repetitive lifting from ground level. The next three slides demonstrate the lifts

Jan 1, 1986

ITEC Environmental Colorado has begun reprocessing the tailings from old mining sites near Boulder, CO. And if the technology behind ITEC&apos;s pilot project at Gold Hill proves successful and profitable, the company plans to launch similar reprocessing operations in other parts of the state. The pilot project is under the watchful eye of state and federal authorities. They hope the technology becomes the answer to cleaning up old mining sites throughout the West. Remining and environmental cleanup on an economic basis, then, are ITEC&apos;s hopes in one the most environmentally conscious counties in the state. Clean up the mine dump sites and recycle the land, for a profit. In June 1998, ITEC Mineral Inc., ITEC Environmental Colorado&apos;s Canadian parent, acquired the Gold [ ] Hill Mill facility in Boulder County. Included in the acquisition was 4.4 hm2 (11 acres) containing the Gold Hill Mill and 283 hm2 (700 acres) of additional properties adjacent to the facility. ITEC estimates that there are about 910 kt (1 million st) of dump rock in the area. That is enough to make the operation viable. ITEC will sell the metals recovered through its reprocessing operations. However, due to low metals prices, ITEC is not expecting to make much money on the recovery of precious and base metals. Instead, the company plans to resell the reclaimed land for residential and commercial development or open space. The Gold Hill Mill site is the first of 17 sites in Colorado that ITEC Mineral hopes to clean up and recycle the land. Most notable among the sites is the historic Central City/Black Hawk mining district, which ITEC believes holds considerable potential. The Gold Hill Mill is located 14 km (9 miles) west of the city of Boulder and less than a half mile from the small town of Gold Hill. The mill was. last operated in 1992. The operation was forced to close that year due to environmental and cash flow problems. Gold Hill is in the middle of a region with more than 32 Mt (35 million st) of mine tailings. Independent geologists have&apos; established the average grade of the Gold Hill mine tailings to be 4.5 to 5.3 g/t (0.13 to 0.155 oz/st) of gold and 28 to 44 g/t (0.83 to 1.29 oz/st) of silver. According to these studies, the Gold Hill mine tailings are five times richer in precious metals than the tailings of the Norebec Manitou property in Canada that ITEC is also remining, reclaiming and restoring. Upon acquiring the mill, ITEC undertook a refurbishing program that included adding a second ball mill. This increased the Gold Hill Mill&apos;s capacity to 180 t/d (200 stpd). The operation employs 15 people. "The Gold Hill Mill site will demonstrate the efficiency of our integrated remining, reclaiming and restoration process," said Maxine Stewart, president of Denver, CO -based ITEC Environmental. "The abundance of similar mine sites in Colorado presents a number of opportunities to bring the benefits of the ITEC system to other communities around the state." Holly Huyck is technical advisor to the Clear Creek Watershed Advisory Group. She said: "ITEC is cleaning up an orphan mine site that otherwise would continue to contaminate the Boulder Creek watershed with acid and excess metals. This is a win-win situation: ITEC gains marketable minerals while it reclaims mined land and helps to improve the watershed."

Jan 1, 1998

By G. E. Karantzavelos

Introduction Pure alunite (KAl3 (SO4 )3 (OH)6 ) contains 37% A1203. In the alunite ore found on Milos Island, Greece, opal is the principal gangue while some aluminum, iron, and titanium oxides are also present. Processing of alunite ore leads to the recovery of potassium sulfate, alusil (alumina-silica mixture), and sulfur dioxide. An interesting process has been developed in Greece (Svoronos, 1977) consisting mainly of roasting the alunite ore mixed with KCl to disrupt the mineral lattice in preparation for separating its components. The thermal decomposition of alunite at approximately 600°C occurs by the reaction: KAl3 (SO4)3 (OH)6 + nSiO2 + 6KC1 - 4 K2SO4 + 3 A1203 . nSiO2 + 6 HCl + 3H20 According to the process pure alunite mixed with 543 kg KCl, 840 kg K2SO4, 370 kg alusil, 264 kg HCl, and 65 kg H2O are produced. The roasted calcine is leached in hot water. The potassium sulfate leach slurry is then thickened and washed to obtain a solid residue (alusil) plus a solution rich in potassium sulfate. Assuming the alunite ore feed consists of 50% to 60% alunite, the solid residue might consist of approximately 22% to 27% A1203, silica, iron, and titanium oxides. This leach residue could be digested in a hot sodium hydroxide solution to dissolve alumina. An amount of the alumina (approximately 10%) would not dissolve. In addition, part of the silica would dissolve requiring removal from the solution in a subsequent desilication operation. The same leach residue could be attacked by mineral acids such as a diluted hydrochloric acid to extract aluminum chloride (Sawyer et al., 1983). Alumina and aluminum are presently produced from bauxite of which Greece has important reserves. If this aluminous material (alusil) could be beneficiated by any physical process, then it would be considered as a raw material for the production of fire-clay refractories instead of an alternate source of alumina. Experimental procedure Extensive laboratory work (mineralogical and chemical analyses) indicates that alusil is a mixture of liberated (mostly) alumina and silica particles, not a natural silicate mineral. Table 1 shows the chemical assays of two alusil samples that were obtained from a roasted alunite ore leaching pilot plant. The samples used for the experiments were a very fine alusil powder (K80 = 400 mesh, 38 µm). These were pulped in tap water. Desliming of the pulp before each flotation test was ineffective. Flotation tests were performed in a 2000 ml Denver cell. A standard 500-g sample was first conditioned at a pulp density 50% solids for 5 min with hydrofluoric (HF) acid solution (40%). The temperature of the conditioned pulp was measured and kept nearly constant at 50°C. The pH of the conditioned pulp was also regulated to 2.5-3 with HF solution. Then ARMAC C solution (2%) as well as two drops of fuel oil and one drop of AEROFROTH 65 were added and the pulp was reconditioned for 2 min. After conditioning, the viscous pulp was transfered to the flotation cell. It was then diluted with warm water to about 15% solids and floated to separate silica from alumina particles. A rougher froth product that was cleaned and recleaned was obtained from the flotation tests.

Jan 1, 1987

The Sandvik Group (Sandvik AB), headquartered in Sandviken, Sweden, continues to expand its worldwide operations growth. After recent acquisitions, including Tamrock, headquartered in Tampere, Finland, the company now has annual sales of more than US$5 billion (SEK 40 billion). With the acquisition of Tamrock, Sandvik has further increased its interests in the mining industry. Sandvik Rock Tools, a familiar name in the mining industry, is one of the world&apos;s leading suppliers of rock-drilling tools. And Tamrock is one of the world&apos;s leading suppliers of drilling machinery, loaders and equipment for rock-excavation operations. Sandvik believes that the two companies combined offer customers "maximum productivity through total system solutions, including machinery, service and tools." In 1997, Tamrock posted sales of US$940 million (SEK 7.5 billion). In addition to its expertise in drilling tools and equipment, Sandvik is supported by its collective resources in specialty steel and cemented carbide. These resources include extensive research and development capabilities, with annual investments in R&D now at about US$150 million. As a result of the R&D programs, internal expansion and acquisitions, the company has nearly doubled in size during the past four years. From 1994 to 1997, nearly US$1 billion was invested in internal growth and another US$825 million has gone towards acquisitions. Return on net assets is now 22% annually. In addition to Tamrock, acquisitions of interest to the mining community include Drillmaster Inc., Houston, TX. The acquisition of Drillmaster (by Sandvik Rock Tools) has further strengthened Sandvik&apos;s position in down-the-hole drilling tools. Other major recent acquisitions by Sandvik include Kanthal, a Swedish manu¬facturer of metal and ceramic-resistant materials for electric heating furnaces, household appliances and elec¬tronics, and Precision Twist Drill, Crystal Lake, IL, one of the leading producers of high-speed cutting tools. With more than 38,000 employees worldwide, the Sandvik Group is divided into six separate business areas. Each area is responsible for its own R&D, production and sales: ? Sandvik Mining and Construction. Including Tamrock and Sandvik Rock Tools, this business area boasts of offering the most comprehensive range of "rock-working" products. These include drilling rigs, loaders, rock-excavation equipment and rock drilling tools. ? Sandvik Specialty Steel. This area includes Sandvik Steel and Kanthal, manufacturers of stainless steel tube, strip wire and bar, as well as high-alloy steel and specialty metals. ? Sandvik Tooling. This business area consists of tools and tooling for metalworking applications. Sandvik Coromant is a market leader in the development, manufacture and marketing of tools using cemented carbide, ceramics and extremely hard materials, such as diamond and cubic nitrite. CTT Tools is the world&apos;s largest manufacturer of high-speed steel tools and is now expanding operations into solid cemented-carbide tools. ? Sandvik Hard Materials. This segment produces a full range of "hard materials," ranging from cemented carbide to diamond. This area manufactures and markets a full range of cemented carbide blanks for tool manufacturers, as well as components and rolls. ? Sandvik Saws and Tools. This business area manufactures high-quality handsaws and metal saw blades, as well as hand tools ranging from wrenches to chains for logging machines. ? Sandvik Process Systems. This segment manufactures steel belts, process systems and automatic sorting systems for the chemical

Jan 1, 1998

By J. D. Morgan

Demand for nonfuel minerals depends in large measure on economic activity, which rose in each quarter of 1985. However, despite 1.7 million new housing starts and a 7% increase in motor vehicle production, many domestic metal mining operations experienced continued difficulties as a result of import competition, greater use of newer engineered materials, emphasis on multi-unit housing, and downsizing of motor vehicles. Recognizing the continued difficulties of the domestic metal mining industry, many corporations made major efforts to retrench and rationalize - selling, closing, or spinning off less profitable operations, renegotiating labor contracts, and inviting increased participation by foreign firms. About 15% of US energy was imported in 1985. This included $52 billion worth of crude oil and refined products, adding significantly to the overall merchandise trade deficit of $149 billion. With the exception of a few large-bulk, low-value raw materials, most minerals move relatively freely in world trade. The strong dollar encouraged imports and made exporting more difficult US nonfuel minerals trade deficit totalled $17 billion. The US continued to rely on imports for many important strategic minerals in 1985 (see accompanying Table). Increased pressures for protection included major mineral products such as steel, ferroalloys, and copper. But President Reagan, in his Feb. 1986 "Economic Report to the Congress" stated: "Our international trade policy rests firmly on the foundation of free and open markets." He also stressed the need for fair trade and more open markets internationally. The National Critical Materials Council, created by law in the latter part of 1984, was activated in the latter part of 1985. Its first public hearing was held in January 1986. The Council is to review the nation&apos;s materials posture and advise the President on such matters. Compared with 1984, the value of US raw nonfuel mineral output rose 2%, to $24 billion. This supported the production of $244 billion worth of processed materials of mineral origin. Raw steel production - a major indicator of both material use and mineral production - fell 4% to 81 Mt (89 million st). Consequently, domestic iron ore production fell to 49 Mt (55 million st), about 97% pelletized. Iron ore imports fell 7% to 16 Mt (18 million st). Raw steelmaking capacity fell 4% to 117 Mt (129 million st). Pacts with 16 leading steel exporting nations have been negotiated in an effort to reduce import penetration of the domestic market. Domestic producers of bulk ferroalloys of chromium, manganese, and silicon operated at about 40% capacity. The US mined no manganese or chromium. Only a small quantity of ferronickel was produced from domestic laterite mining in Oregon. Molybdenum production increased slightly to 49 kt (54,000 st). Mine shipments of tungsten, another mineral tied to steel production, also fell slightly to 1.1 kt (1200 st) of contained metal. Even the domestic aluminum smelting industry suffered reverses as high stocks and low prices reduced production 16% to 3.5 Mt (3.8 million st). Similarly, imports of bauxite fell 7% and alumina 10%. Production of magnesium metal fell 6% to 136 kt (150,000 st), while titanium metal production remained unchanged at 22 kt (24,000 st). Copper mine production was unchanged at 1.1 Mt (1.2 million st). Copper refinery production fell slightly to 1.1 Mt (1.2 million st). Lead mine production rose 24% to 400 kt (440,000 st). Zinc mine output was down 10%, to 225 kt (250,000 st). Refined lead production increased 28%, to 510 kt (560,000 st). Primary slab zinc production fell 5% to 240 kt (260,000 st).

Jan 5, 1986

By Nathaniel Arbiter

The development of flotation machines since the first industrial use of the process can be divided into three stages. The first, up to about 1930, was marked by diversity in the means for contacting floatable mineral with the air/ water interface and for generating the interface, as well as in cell design. By 1927 Taggart (1) could list 170 patents covering flotation cells, and he mentioned dozens which were or had been in operation. He divided cells into eight categories according to design, type of separating interface, and where in the cell and how separation took place. These were: (1) Film or skin flotation with wet or dry oiling of surfaces; (2) pulp body process with bubbles generated chemically by pressure reduction, by vacuum or by boiling; (3) agitation-froth processes involving local supersaturation with air; (4) bubble column; (5) pneumatic; (6) cascade; (7) subaeration; and (8) combination and miscellaneous types and processes. Most widely used then were the agitation-froth, sub- aeration, pneumatic types. By 1943 a significant reduction in categories had occurred to the two main divisions of mechanical and pneumatic cells according to whether agitation and aeration were (a) by an impeller with or without compressed air, or (b) by compressed air alone. Cells at that time were still relatively small; the Morenci Concentrator completed in 1942 used 432 1.7m3 (60 cu ft) Fagergrens for roughing 45,000 and later up to 55,000 t/d.

Jan 1, 1979

By Larry F. Barrett

Anaconda exploration in the Rico mining, district from 1978 through early 1983 resulted in discovery and partial definition of the Silver Creek molybdenum deposit. Approximately 44 million tons of 0.31% Mo are drill-indicated, and projections suggest that the deposit exceeds 200 million tons. No source intrusion has been intersected by drilling, but its composition is suggested by silicic alkali-alaskite porphyry dikes of intramineral age observed at the surface and in drill holes. Age data and fluid inclusion studies indicate that the deposit was formed 5.2 ±2 m.y. ago and that more than 1,200 meters of cover has since been removed by erosion. Both emplacement of the hypothesized source intrusion, and the geometry of the molybdenum deposit, were influenced by three major intersecting faults. The most significant of these, the Last Chance fault, juxtaposes Precambrian quartzite and greenstone against Pennsylvanian clastic and calcareous sediments within the mineralized zone. Wallrock alteration associated with the deposit is characterized by potassic, phyllic, and propylitic zones in non-calcareous rocks, and garnet and anhydrite-diopside zones in the carbonates. All +0.2% Mo mineralization is within the potassic and garnet zones. Pre-mineral faults that intersect the deposit contain the only surficial molybdenum, tungsten, and fluorine anomalies. Pervasive dispersion haloes of the indicator elements were intersected at depth by drilling in the discovery phase of exploration, and are particularly well-developed in the hanging wall of the Last Chance fault.

Jan 1, 1985

By E. W. Lindroos, Fred D. DeVaney, R. L. Bennett, R. W. Hernlund, W. R. van Slyke, W. F. McDermott, C. U. Keranen

INTRODUCTION Iron is the world&apos;s most widely used metal and constitutes about 5% of the earth&apos;s crust. Iron ores are found widely distributed throughout the world. Because of its wide occurrence and relatively low value, an iron ore to be merchantable must contain a relatively high percentage of the metal to be considered as a source of iron. Few ores containing less than 25% iron would be considered as ore unless they existed in large amounts and could be concentrated very cheaply. The grade of the ore, the ease of concentration, and the transportation cost to market are fundamental considerations deter- mining its worth. Iron never occurs as a natural metal except in meteorites. Most iron ores occur as oxides, and to a lesser degree as carbonates. A small amount is recovered as a byproduct from iron sulfides. Iron ore must be reduced (oxygen removed) to metal for use. Iron is marketed in many forms and combinations, such as pig iron and a great variety of steels. Iron is practically never pure as marketed but contains varying amounts of carbon, phosphorus, sulfur, and alloying elements such as nickel, chromium, etc., all of which affect the physical proper- ties of the end product. Please note that in this section all analyses ore on a dry basis and all tonnage figures are on a long-ton basis of 2,240 lb unless otherwise noted.

Jan 1, 1985

By C. H. Bucknam

Acid neutralization titrations were applied to the characterization of soil and overburden for the control and prediction of acid rock drainage in the coal fields of the eastern United States in the 1970?s (1). These became known as the Sobek methods as a result of the early United States Environmental Protection Agency publication and the methods have been modified (2) to be adapted to many applications in the metal mining sector (3,4,5,6,), and, as a result, there are a large number of acid neutralization potential methods being used that are referred to as Modified Sobek methods. Standardization of neutralization potential methods for metal bearing ores and related materials based on carbonate carbon, rather than titration, commenced in the 1990?s through the American Society of Methods and Materials (ASTM) Committee E01 on Analytical Chemistry for Metal Ores and Related Materials Subcommittee E01.02 on Ores, Concentrates and Related Metallurgical Materials (7). A subcommittee section, E01.02.05 on Environmental Methods, was organized to coordinate the method and material standardization efforts of ASTM and the Acid Drainage Technology Initiative Metal Mining Sector (ADTI-MMS) Prediction Committee (8), which took on the project of standardization of the acid neutralization potential acidity titration method (ANPA) to supplement the carbonate carbon method in E1915 (9) for application to samples which have metal carbonates (10), in particular, which result in overestimation of the neutralization potential using the carbonate carbon methods (11).

Jan 1, 2010

By R. D. Rein

The Henderson mine is a large continuous panel caving operation in central Colorado. The orebody is roughly 760 m by 910 m and will be extracted from two production levels. Each level is divided into three north - south panels 244 m to 317 m wide. The general arrangement of a production level and its associated undercut drifts is shown in Figure 1. The production drifts (3.6 m by 3.7 m) are spaced 24 m apart center to center and have undercut drifts (3.4 m by 3.8 m) 17 m directly above them. The production drifts are interconnected with drawpoint crosscuts (4 m by 4.3 m) driven on 12 m centers. Drawpoint crosscuts receive substantial brow steel and concrete support prior to undercutting. The brow steel is a 15 cm by 46 cm wide flange beam with a 2.5 cm wear plate to box in the &apos;cave&apos; side of the beam. The concrete is a minimum of 0.3 m thick and the batch design results in 27 500 kPa strength. The finished drawpoint opening is 3.7 m by 3.2 m with a choke area of 3.7 m by 2.4 m. (Figure la) Undercutting and bell formation is accomplished by blasting rings of 7.6 cm holes drilled from the undercut drifts. The rings have a burden of 2.0 m and six rings develop one complete bell.

Jan 1, 1985