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By V. O. Tenorio, A. M. Sanzana

"INTRODUCTION This report summarizes a research about mines which are transitioning from surface to underground mines. It is a guide to make strategic decision in these increasing amount of mines. Capital and operating costs are linked with major strategic decisions in the exploitation of an ore body. This is the reason to analyze costs related to this transitioning world trend. There are multiple reasons for transitioning from Open-Pit (OP) to Underground (UG) mining. The primary economic incentive for going UG is that the cost per unit of ore from OP mining eventually exceeds that from UG extraction. This is due to the excessive amount of waste that must be extracted when the depth of the pit increases. In addition, the reduced footprint of UG operations contributes to diminish the environmental impact on the surface. Furthermore, it becomes more difficult to find new, easily accessible ore deposits. Then, a cost model helps to guide strategic decisions in the transition from OP-to-UG mines. Objectives • Propose a methodology to guide strategic decisions in the transition from OP to UG mining. • Create an integrated cost model based on existing OP and UG individual cases. • Identify and characterize a group of mines transitioning from OP to UG. • Analyze operating and capital costs of a particular transitioning from OP to UG mine. MAIN CONCEPTS Open Pit Mining (OP) The main decisions in OP mines are: Cut-off grade, Stripping Ratio (SR), equipment size, stockpiling, pushback sequencing and processing capacity. The stripping ratio is measured as the tonnage of waste required to be removed to reach and extract one tonne of ore. The material extracted from an OP exploitation can be classified as waste and ore. This classification depends on the cut-off grade. The cut-off grade indicates the minimum amount of material with economic value in a block. The waste from the mine is going to the waste dump and the ore is going to stockpiles where will be store and then going to the processing plant (leaching or concentrator). The waste of this process is going to the tailing pond big pools. And the product is sell in the market. Another important issue is the mining capacity which is going to define the investment in mining equipment. Underground Mining (UG) The first decision to be made in order to open an underground mine is selecting the mining method. Caving methods are the more productive and have lower operating costs. There are three caving methods: Block Caving (BC), Panel Caving (BC) and Sublevel Caving (BC). The access development could be an horizontal adit or a straight or downspiral decline. A shaft could be used for inclined or vertical access. Other key issues are: equipment capacity, cut-off grade and production sequencing."

Jan 1, 2016

By A. Redman

Aquatic organisms are sensitive to free cyanide (HCN) though the presence of certain metals that can form stable complexes and thereby reducing both the amount of bioavailable cyanide. Recent work has suggested that consideration of HCN rather than total cyanide can provide the basis for water quality criteria. Cyanide is often present in complex effluents that affect speciation and bioavailability. We have developed a model to predict the speciation of cyanide in complex effluents and have shown that this model can predict the toxicity of metal-cyanide complexes in terms of HCN. This study evaluated available cyanide toxicity exposures with this model to assess the role of water chemistry (e.g., metal-cyanide interactions) on the observed. Toxicity endpoints based on total cyanide ranged over several orders of magnitude for a variety of metal-cyanide mixtures. However, predicted HCN concentrations among these same tests described the observed toxicity data to within a factor of 2. Aquatic toxicity can be well described using HCN though in certain metal-cyanide mixtures toxicity was jointly described by HCN and elevated levels of bioavailable metals. This work was performed in support of an industry-sponsored reassessment (Gensemer et al 2006) of the U.S. EPA water quality criteria for cyanide (EPA 1985) using free cyanide as the basis for the re-calculation. The work presented in this manuscript presents technical support for using free cyanide as the preferred basis for this recalculation.

Jan 1, 2011

By G. Robson

"From monitoring gas levels and air quality in coal mines to protection from collisions for large vehicles at mine sites, SICK offers a wide range of sensors and systems to improve safety and to monitor potentially dangerous situations. SICK offers solutions for the analysis and monitoring of underground coal mining atmospheres. These systems supply the coal mine operator valuable safety and ventilation information, as well as data which identify and monitors atmospheres of those production areas within the mine which may or may not be worked, such as Goafs and sealed areas. Underground coal mines contain dangerous gases, which must be monitored and controlled in order to allow productive mining operations to take place and to minimize the risk of explosion. Underground mines are a potentially dangerous place for mine workers. These are the main reasons for using Tube bundle systems SICK has had great success in selling these systems into the Australian and New Zealand Markets and this is helped by the strong drive for mine safety from both the Mine Owners and the Regulatory bodies charged with the responsibility of ensuring Mine Regulations are met. The Australian mining industry uses Tube Bundle Systems at almost all underground sites, even after mining has ceases and the mine is placed in care and maintenance. Sick has been working in international markets for some time now but has had limited success in selling Tube bundle Systems to these markets. However in recent times there has been raised interest in Tube Bundle from China, India and Turkey as these countries strive to improve mine productivity and lessen the risk of explosion or Spontaneous Combustion. Our lack of penetration into International Coal markets has been very much based on the perceived cost of the system, not the benefits to safety and productivity. Unless a mine has had an event such as spon com or a Methane explosion they rarely consider the very low level measurement capability (20 PPM CO & 100 PPM CH4) or the long term trending benefits of the Tube Bundle system. It is only after a catastrophic event the System is considered. Real Time or electrochemical Sensors are commonly used to monitor the Mine Atmosphere, the real problems of methane build up or potential Spon Com in Goafs and sealed areas cannot be monitored using Real Time Sensors as the electrochemical cell deteriorates after approx. 12 months and in sealed areas these cells cannot be maintained or replaced. Whilst the Technique of Tube Bundle Systems is quite old SICK have spent many years redesigning the Tube Bundle System into a cost effective and real value for money product. The system uses state of the art sampling and highly accurate Infra-Red wide Spectrum Analysers."

Jan 1, 2016

By D. M. Avant, A. J. Solomon, S. M. Pickering, W. R. Fox

Introduction Most high purity silica is produced from crystalline quartz, generally from sand. However, there is a small area of southern Illinois where a porous, friable white chert rock is found that is very different in physical form from quartz sand. This microcrystalline material, commonly referred to as tripoli, is more easily ground to micron sizes than sand. Although of similar Moh's hardness, microcrystalline tripoli is less abrasive than either ground quartz sand or many synthetic abrasives. Location The southwestern corner of Illinois is underlain by extensive deposits of very porous chert; in some places the material is quite white. Two companies, Illinois Minerals Co. and Tammsco Inc., mine and process this white material. The Illinois Minerals plant is located at Elco, and the Tammsco plant is in Tamms, IL. Both companies mine from various locations within a 10-mile truck haul of their plants. Production from this district has been continuous since at least 1915. Malvern Minerals Co. mines and processes a generally similar microcrystalline, friable novaculite near Hot Springs, AR. Geology The southern Illinois white tripoli area is on the southern flank of the Illinois Basin. The commercial silica zone is a facies of the Clear Creek chert of Lower Devonian Age. To the north, where it occurs deep in the basin, the Clear Creek is a siliceous limestone. In the mining area, thin overburden has allowed access by surface ground water, which has leached away the carbonate content and left a quite pure silica residue. Fossils characteristic of sea bottom deposition, such as brachiopods and trilobites, are occasionally encountered in mining. The structure of the mining area is poorly known, due to infrequent outcrops. Strata in most mines are inclined at dips of 5° to 20°. There appear to be occasional high angle normal and reverse faults, but these are seldom of any significance in mining. The friable white silica zone mined in the Clear Creek chert is overlain by dark brownish-red chert and siliceous shale. Exploration drilling to date has not encountered the base of the microcrystalline tripolitic zone. Chemical content Leaching by ground water solution has removed virtually all the carbonate from the original siliceous limestone. Impurities encountered are red to yellow iron oxides, aluminous clays, and occasional black manganese oxides. The silica has no detectable formula water. A typical chemical analysis and physical properties are listed in Tables 1 and 2. Mineralogy and crystallinity Southern Illinois friable tripoli appears amorphous under the polarizing microscope. No crystal structure and only occasional, vague, patchy anisotropism is evident. However, the material gives an excellent quartz X-ray diffraction pattern. Thus, so far as X-ray mineralogy is concerned, the material is clearly

Jan 12, 1986

By M. J. Gregory

The Pebble Cu-Au-Mo deposit in southwest Alaska is one of the world's largest porphyry deposits. The western part of the deposit was discovered by Cominco America in 1987, who sold the project to Northern Dynasty Minerals Ltd., in 2001. Northern Dynasty discovered the high-grade eastern part of the deposit in 2004 and went into a 50:50 partnership with Anglo American PLC in 2007, forming the Pebble Limited Partnership (Lang et al., 2008). The Pebble Limited Partnership has continued to delineate the deposit through diamond drilling with the most recent resource estimate released in February 2010 (Pebble Limited Partnership, 2010). The Pebble deposit mineral resources comprise 5.94 billion tonnes of measured and indicated resources grading 0.78% Cu EQ, containing 55 billion pounds of copper, 67 million ounces of gold and 3.3 billion pounds of molybdenum. Also delineated are 4.84 billion tonnes of inferred mineral resources grading 0.53% Cu EQ, containing 25.6 billion pounds of copper, 40.4 million ounces of gold and 2.3 billion pounds of molybdenum (Pebble Limited Partnership, 2010). Resources are calculated using a 0.3% Cu EQ cutoff. GEOLOGY AND MINERALIZATION Deposit Geology The Pebble deposit is located in the Kahiltna terrane, between terminal strands of the north-east-striking Lake Clarke strike-slip fault. This fault represents an accretionary boundary and separates the Kahiltna terrane to the north-west from the Peninsular terrane to the south-east (Lang et al., 2008). The Pebble deposit represents a large magmatic-hydrothermal system centered on a group of 90 million year old porphyritic hornblende granodiorite intrusions. An east-west cross section through the southern part of the deposit (Figure 1A) shows the typical relationships between the various rock units which host sulfide mineralization. Sub-parallel ~96 million year old granodiorite and diorite sills were intruded into siltstones of the Jura-Cretaceous Kahiltna flysch. In the west, these sills were cut by a ~96 million year old alkalic intrusive complex composed of porphyritic monzonites and monzodiorites and related intrusion breccias (Bouley et al, 1995). The mineralized 90 million year old hornblende granodiorite intrusions cut all rock types and comprise four small stocks in the western part of the deposit and a larger pluton in the eastern part of the deposit. Recent deep drilling indicates that these intrusions coalesce at depth. The strongest porphyry-style mineralization is centered on the eastern granodiorite pluton and extends into the surrounding rock units (Figure 1A). The western half of the deposit crops out at surface, whereas the eastern side is un-conformably overlain by a cover sequence of Late Cretaceous to Eocene sedimentary and volcanic strata. The deposit was tilted 20º to the east during the Eocene.

Jan 1, 2012

By Gary Krix

The Amdel Thickener Interface Gauge (TIG) provides continuous on-line measurement of the level of the liquid/solid interface in thickeners and Counter Current Decantation (CCD) washers. Unlike other interface level measurement systems, the TIG also determines the span of the transition zone. The transition zone is the region between the clear liquid overflow zone and the compressed solids zone. The Thickener Interface Gauge determines the level of the interface by measuring the natural radioactivity of the solids in the thickener tank using an array of up to twelve detectors mounted vertically in a stainless steel probe. All solids contain low levels of natural radioactivity while the liquid phase generally does not. The concentration of natural radioactivity at each level in the thickener therefore depends on the concentration of solids. The radiation emitted by the solids is converted to electrical pulses by the detectors in the probe and the pulse-rate provides an accurate measurement of the density. A typical count-rate profile is presented in Figure 1. The vertical axis indicates the depth in the tank referenced to the surface of the liquid. The horizontal axis indicates the count-rate measured by the TIG probe, this count-rate is proportional to the density and can be used to determine the settling rate. The count-rate at the top of the graph is relatively low, this region corresponds to the clear overflow zone which contains virtually no solids. Moving down the vertical axis the count-rate increases dramatically at the level of the transition zone where the solids content increases with the depth in the tank. In this region the settling conditions of the solid particles change from free settling to compression. Moving further down the vertical axis the count-rate reaches a plateau where the count-rate remains relatively constant although there are often small fluctuations, this region corresponds to the solids compression zone. The liquid/solid interface as measured by the TIG corresponds to the depth at which the count-rate crosses the target count-rate which is generally selected to be midway between the liquid count-rate and the compressed solids count-rate. The length of the TIG probe is selected to cover the typical operating range of the liquid/solid interface in the thickener or washer. If, for example, the minimum bed level is 3.0 metres below the surface of the liquid then the probe length is selected to extend 30 metres below the surface. The TIG probe can include up to twelve detectors and the se are spaced over the measurement range. For small measurement ranges of around one metre 4 to 6 detectors would normally be used while for larger measurement ranges of three metres or more twelve detectors would be supplied.

Jan 1, 1993

By Jurgen H. L. Stuve, Rick E. M. Widelski

Any one who has been involved with the mining or the treatment of gold-bearing ores will agree that at the end of the total process, the quantity of gold recovered is never as much as was expected. These short falls are usually explained away by claiming that:- the original evaluation of the ore being mined was inaccurate the plant treatment process is not at its optimum theft is taking place excessive dilution occurs during mining The truth of the matter is that the gold is being "lost". During all the stages of its production, starting from the mining, right through to the smelting process, a certain amount of gold is not recovered. When a newly found gold deposit is being evaluated and the drilling data being analysed, cautionary factors are applied so that the grade is not over-stated. Once the decision has been taken to start mining and development is in progress, the exposed ore is sampled. When Again, when this data is being analysed more safety factors are used to ensure that when a piece of ground is given a block-value, the gold content is not over estimated. mining actually takes place and the ore transported to the plant for treatment, the metallurgical department also introduces certain safety factors when predicting how much gold will be recovered. This shows that the quantity of gold that one expects to finally recover is already a conservative figure. The actual quantity recovered invariably falls short of this figure, in fact, South African gold mines are satisfied when achieving a Mine Call Factor (MCF) of 85%. In the last few years a great deal of work has been carried out to investigate how this unaccounted for gold is lost and what steps can be taken to remedy the situation. Previously mined areas were re-swept using the traditional hand brush method; experiments were carried out with water jets sometimes in conjunction with compressed air and even small portable vacuum cleaners were tried. None of these methods were totally successful in recovering the left over gold but the use of vacuum systems showed the most potential. the In South Africa this has resulted in introduction of high-powered vacuum systems on the gold mines to suck-up this lost gold. These were initially introduced in the mining areas, particularly old areas, to recover any left over ore. The results proved extremely profitable, So much SO that presently, well over 120 suction systems are doing this type of work. This success encouraged plant operators to look at areas of potential loss in the plant itself. This has resulted in an increasing number of machines being supplied to gold plants mainly for clean-up applications with the bonus that the material sucked up invariably contains gold.

Jan 1, 1993

By Vas P. Kenyen

Introduction Autogenous, semi-autogenous, and conventional grinding circuits are used to process ores. Current emphasis, though, is towards the use of the first two systems. The reason is the generally lower initial investment required for an autogenous or semi-autogenous circuit as compared to a conventional comminution circuit. However, the effect of operating costs that will significantly affect overall comminution economics is not considered in this evaluation. About 70% of the energy in a mineral processing plant is used in the comminution circuit. Until the 1974 oil embargo, energy was relatively inexpensive. Since that time, the cost of energy has escalated 40-fold. As a result, energy cost now forms a significant portion of operating costs in comminution. According to a survey conducted by General Electric Co. (Powers, 1982), the price of electricity will continue to rise 15% to 20% per year, in spite of the current oil glut. Both autogenous and semi-autogenous grinding circuits consume more energy than the conventional comminution circuit. Higher consumption ranging from 16% to 35% has been reported in the literature (Bassarear, 1982; Gjerdrum. 1979; Betts, 1982; and Christie, 1979). Consequently, even a small increase in energy costs is likely to have a significant impact on operating costs. Therefore, the economic decision of selecting an optimal comminution circuit is important. More than ever before, it is now critical to include in the evaluation of communition circuits not only the capital costs, but also the operating costs; in particular, the expected energy cost over the life of the operation. Failure to do this may lead to a less than optimal comminution circuit. Design A comminution system should be designed only after bench-scale or pilot plant testing of a representative sample. Conventional crushing and grinding circuits can be designed on the basis of the Bond Work Index (Rowland and Kjos, 1980). Over the years, sizing of full-scale rod and ball mills from Bond Work Index data has become well established. Autogenous and semi-autogenous mills cannot be sized based on bench-scale tests. Pilot plant tests are necessary to determine the amenability of the ore, critical size buildup, and power requirements. For this study, conventional and semi-autogenous grinding circuits were designed for ore throughput of 12, 24, 36, 48, 72, and 90 kt/d (13,200, 26,400, 39,600, 52,000, 79,200, and 99,000 stpd). Rod and ball mills were scaled up on the basis of the Bond Work Index. A two-stage semi-autogenous grinding circuit was designed on the basis of pilot plant test results. In designing a grinding system, emphasis was placed on operating flexibility. The largest commercially available equipment was

Jan 10, 1984

By D. M. Avant, A. J. Solomon, S. M. Pickering, W. R. Fox

Introduction Most high purity silica is produced from crystalline quartz, generally from sand. However, there is a small area of southern Illinois where a porous, friable white chert rock is found that is very different in physical form from quartz sand. This microcrystalline material, commonly referred to as tripoli, is more easily ground to micron sizes than sand. Although of similar Moh's hardness, microcrystalline tripoli is less abrasive than either ground quartz sand or many synthetic abrasives. Location The southwestern corner of Illinois is underlain by extensive deposits of very porous chert; in some places the material is quite white. Two companies, Illinois Minerals Co. and Tammsco Inc., mine and process this white material. The Illinois Minerals plant is located at Elco, and the Tammsco plant is in Tamms, IL. Both companies mine from various locations within a 10-mile truck haul of their plants. Production from this district has been continuous since at least 1915. Malvern Minerals Co. mines and processes a generally similar microcrystalline, friable novaculite near Hot Springs, AR. Geology The southern Illinois white tripoli area is on the southern flank of the Illinois Basin. The commercial silica zone is a facies of the Clear Creek chert of Lower Devonian Age. To the north, where it occurs deep in the basin, the Clear Creek is a siliceous limestone. In the mining area, thin overburden has allowed access by surface ground water, which has leached away the carbonate content and left a quite pure silica residue. Fossils characteristic of sea bottom deposition, such as brachiopods and trilobites, are occasionally encountered in mining. The structure of the mining area is poorly known, due to infrequent outcrops. Strata in most mines are inclined at dips of 5° to 20°. There appear to be occasional high angle normal and reverse faults, but these are seldom of any significance in mining. The friable white silica zone mined in the Clear Creek chert is overlain by dark brownish-red chert and siliceous shale. Exploration drilling to date has not encountered the base of the microcrystalline tripolitic zone. Chemical content Leaching by ground water solution has removed virtually all the carbonate from the original siliceous limestone. Impurities encountered are red to yellow iron oxides, aluminous clays, and occasional black manganese oxides. The silica has no detectable formula water. A typical chemical analysis and physical properties are listed in Tables 1 and 2. Mineralogy and crystallinity Southern Illinois friable tripoli appears amorphous under the polarizing microscope. No crystal structure and only occasional, vague, patchy anisotropism is evident. However, the material gives an excellent quartz X-ray diffraction pattern. Thus, so far as X-ray mineralogy is concerned, the material is clearly

Jan 1, 1987

JIM ARNOLD: Questions? Comments? Gary, were you gone when I made the comments about corporate puke? [laughter] SPEAKER: You were hiding. BRUCE CAVENDER, BHP Copper: Having worked in both places, I'd like to address the issue of the plant versus the corporate office. One thing to keep in mind is that we are all very engaged with the technical elegance of a lot of the work we do. There's some tremendous technical expertise in the plant. But remember, the guys in the corporate office are only thinking in one denomination, and that's dollars. So we'll see something like a spill, and for us it's a problem. At the end of the day, what the corporation cares about is the financial impact. So we've got to be careful to couch our solutions in dollar terms. JIM ARNOLD: Rebuttal? [laughter] DOUG HALBE: Are you going to rebut this one, Andy? ANDY CARTER, Oxidor Gold Corporation: No, I don't think so. I'm not up to it. I'd just like to address a question both to our panel and the floor. Those of us who've been in this industry any length of time have seen these same mistakes made over and over again. Wherever you go, year after year, plant after plant, you see the same problems. I guarantee you, Doug, that if we have a similar conference down the road, you're going to see the same thing. My question to the panel and to the floor is what is it about our industry that stops us from learning from our mistakes? What characteristic is it that prevents us from learning and makes us repeat the same mistakes over and over again? DOUG HALBE: That's a great question. Someone said in a meeting 10 years ago that if the people who design plants designed airplanes, the planes would be falling out of the air all the time. GARY SIMMONS, Newmont Gold Company: I've got an answer to your question as to why many plants are poorly designed. It's that the minute anybody demonstrates any sort of competence in plant design, they're promoted to a higher-paying job, and the talent pool is left with people of lesser experience. Just think about it: The people who have the experience and have demonstrated it more than once have moved up. JIM ARNOLD: I think that there's another reason, too. Seven or eight years ago I was chairing a session, and I challenged everyone in the session to some day publish an article about the design mistakes they'd made because we never talk about our mistakes. I'll

Jan 1, 1998

By S. N. Groudev

A process for the combined chemico-biological removal of iron present as oxide minerals in different mineral raw materials (quartz sands, kaolins, clays, etc.) was developed. In the process, mineral raw materials are leached at about 90 °C with lixiviant containing microbially produced oxalic and hydrochloric acids. The leaching is carried out in mechanically stirred acid-resistant reactors for periods of from 1 to 6 hours, depending on the iron content and the forms of iron in the raw materials being leached. The iron contents of some sands treated by this method were lowered from levels that were in the range of 0.035% to 0.088% Fe2O3 to less than 0.012% Fe2O3, making them suitable for the preparation of high-quality glass. The iron contents of different kaolins were lowered from levels that were in the range of 0.65% to 1.49% Fe2O3 to levels in the range of 0.44% to 0.75% Fe2O3. As a result, the whiteness was increased from values of 55% to 87% to values of 86% to 92%. The iron content of clay was lowered from 6.25% Fe2O3 to 1.85% Fe2O3, and this increased the “fireproofness" of the clay from 1,670°C to 1,750°C. A similar process was used for the leaching of aluminum from aluminosilicates, mainly clays and kaolins. However, in this case, the microbial fermentation fluid containing citric acid was acidified by means of sulfuric or hydrochloric acid or by means of different mixtures of mineral acids. For enhancing aluminum solubilization, the aluminosilicates were heated before leaching at 600° to 650°C for 1 to 2 hours. Over 90% of the aluminum present in different clays and kaolins were leached within 3 to 6 hours. "Silicate " bacteria related to the species Bacillus circulans were used to leach silicon from low-grade bauxite ores containing aluminosilicates as impurities. The bacterial action was connected with the degradation of the mineral structures (by means of microbial metabolites such as organic acids and exopolysaccharides), as well as with the selective separation of the rich-in-aluminum fine fractions, which were retained by the mucilaginous capsules of the bacteria. The solid residues after treatment were characterized by higher values of alumina and silicon module (A1203: SiO2 ratio), and they were suitable for processing by means of the Bayer process for recovering aluminum. The bending strength and other ceramic properties of kaolins were improved by contact with well-developed cultures of "silicate" bacteria. The improvement was caused mainly by bacterial metabolites (exopolysaccharides) that acted as resins during drying.

Jan 1, 2000

By James Watson Baxter

Fluxes are used to promote pyrometallurgical processes that rely on adhesion (soldering or brazing) or fusion (gas and arc welding) to join metallic surfaces. In the adhesive processes, the metal surfaces to be joined are not melted; the join is formed using a filler metal with lower melting point than the base metal. Fusion welding involves use of heat in excess of the melting point of the base metal. The fused joint may be achieved either by simply fusing together metal surfaces brought in contact with each other or by introducing additional molten metal of similar composition to form a fused joint. ADHESIVE PROCESSES--SOLDERING AND BRAZING In order for molten filler metal, solder or braze, to spread in a manner that creates a successful join; the work surfaces on the base metal must be thoroughly cleansed. Fluxes remove stubborn oxide films and other surface contaminants, promote wetting of the work surfaces, add fluidity to the solder or braze, and enhance workability and ease of spreading. Brazing processes involve higher temperatures than those reached in soldering. Brazing fluxes, which must remain active and effective at the higher temperatures, differ from those employed in soldering. Some common fluxes used in adhesive processes are rosin for soldering tin and electrical connections, hydrochloric acid for use in soldering galvanized iron and other zinc surfaces, and borax for brazing. Soldering and brazing are similar processes, the primary difference being the temperature at which the joining operation is carried out. Soldered joints, produced with low-melting-point fillers (solders) that melt and flow at temperature less that 450°C (Althouse et al., 1988) can sustain loads of 1 to 1.7 MPa for extended periods of time (Anon., 1966). Brazing involves the use of filler materials with melting points commonly above 500°C and generally provides stronger joints than those obtained with solder. Both processes require local application of heat to melt and spread the filler so that the molten filler can wet (adhere to) the base metals by alloying and diffusion. Soldering Soldering is a means of joining metals by adhesion using a metallic bonding alloy as the filler, commonly a mixture of lead and tin. However, the adhesion of solder depends more on its ability to be keyed into minute surface irregularities than on alloying. The most familiar application is to provide and secure electrical connections. Soft solders can range from 1 to 70% tin with the remainder mostly lead. However, for general-purpose, soft-solder work, the alloy is commonly 50% lead-50% tin. Higher lead contents provide a wider range in the melting temperature and, for this reason, a 60% lead-40% tin alloy, which yields a mushy mixture, is used for wiped joints in lead sheet and pipe work. Conversely, 40% lead-60% tin alloys are used in soldering tin and other low- melting-point materials for which a narrower range of melting temperature is required. There are numerous other solder compositions such as tin-silver, 95% tin-5% silver and antimony-tin, 95% tin and 5% antimony (Carlin, Jr., 1992). Heat needed to melt and spread the solders is commonly provided by electrically heated, copper-tipped soldering irons or by means of torches; the solder is applied by hand, usually face-fed by means of wire. For wiped joints in plumbing and lead-cable splicing, the solder is manipulated with cloth pads. The molten solder wets the joint surfaces and is drawn, by surface tension, into minute fissures and capillary openings. Other applications involve use of induction heaters and furnaces with pre-shaped solder appropriately placed prior to fluxing and heating. In some processes, the joints are immersed in molten solder. Constituents and Role of Soldering Fluxes: Soldering fluxes generally fall into one of three categories: highly corrosive fluxes, intermediate fluxes, and noncorrosive fluxes. These same categories are sometimes designated inorganic, organic and rosin-based respectively (Althouse et al., 1988). Common constituents of each group are discussed briefly below. Corrosive Fluxes (Inorganic). Work with aluminum, magnesium, stainless steel, high alloy steel, aluminum bronzes, and silicon bronzes is carried out at temperatures in the upper portion of the range for solder operations. Soldering these materials requires use of highly active, corrosive fluxes to remove and prevent the formation of the especially stubborn, hard, oxide films that form on these materials upon exposure to the atmosphere. The corrosive fluxes consist of inorganic acids and salts that are applied either as pastes or dry. They are active at elevated temperatures and, since they remain active after the soldering is completed, must be completely removed. The main constituent of most corrosive fluxes is zinc chloride with a melting temperature well above the solidus temperature of most commercial tin-lead solders. It is made by the action of hydrochloric acid on zinc. When zinc chloride is used alone, un- melted particles of this corrosive salt get caught up in the joint and weaken it. For this reason, other inorganic salts such as ammonium chloride (NH4Cl) or sodium chloride (NaCl) may be added to lower the melting temperature. A mixture of zinc chloride and ammonium chloride is very effective because the excellent oxide reducing properties of ammonium chloride and the protective action of the molten zinc chloride combine to produce a fluxing action superior to that achieved when either is used alone. In addition to zinc chloride, ammonium chloride, and sodium chloride; common con-

Jan 1, 1994

By M. C. Stuttle

MDL Rock Lasers has undertaken successful mapping trials of abandoned mine workings using its underground cavity scanning system. The work was being performed for Kalgoorlie Consolidated Gold Mines (KCGM), Australia's largest gold producer. MDL Rock Lasers supplies underground surveying systems to the mining and quarrying industry KCGM's Fimiston Super Pit, located in Kalgoorlie, Western Australia, is Australia's largest open-pit gold operation. Fimiston's final pit design is expected to be 3.7 km (2.3 miles) long, 1.5 km (0.9-miles ) wide and 540 m (1,770 ft) deep. During the early days of Australian mining, underground operations were labor intensive. Mining was selective and generally narrow stopes were mostly backfilled. Improved processing technology and mine mechanization during the 1970s and 1980s has permitted larger scale operations to adopt less selective underground-stope operations. The "super pit" concept was a logical step for mining companies working toward economies of scale. However, the industry recognized that there was no precedent for excavating a 500-m- (1,640-ft-) deep open pit through ground previously affected by old workings. To appreciate the geotechnical challenge facing KCGM, it is necessary to consider that under the pro¬posed area at the open pit, stoping reached a level of 1 km (0.6 miles) and was supported by more than 2,000 km (1,240 miles) of development headings. The company was fortunate because plans were available for the Golden Mile underground operations. The existing plans were interpreted by the mine's geotechnical team and computer models were constructed of the underground workings. These models can be imported into three-dimensional mine planning packages, such as the Vulcan's Unix-based software, Envisage, to assist analysis and design. This prior knowledge of the workings provided the focus for investigating and confirming ground conditions using probe drilling. After several iterations of drilling, followed by detailed analysis, several decisions were made relating to the nature of the ground and the mining approach to be taken. It was recognized that the stope models were not perfect due to hu¬man error, lack of original survey information and progressive deterioration in ground conditions. It also became apparent that the sole reliance on probe drilling was inefficient in terms of time and quality of information. This was particularly true in complex areas where stopes are in close proximity. A major concern was the presence of open stopes (voids). In many cases, stopes consist of combinations of filled and void sections. The condition of pillars within these underground workings is extremely important. In time, pillars collapse and voids will propagate in upward and lateral directions. So systems needed to be developed that allowed rapid and accurate verification of ground conditions. KCGM was introduced to MDL Rock Lasers while investigating technology for underground mining. The two companies decided to use the C-ALS, MDL Rock Laser's underground laser, cavity scanning system, to improve the management of mining through affected ground. This system was to be used in conjunction with probe drilling and void mapping activities. A field test in one area of the Fimiston Mine verified the capability of the C-ALS system to quickly and efficiently assist in this decision making process. The C-ALS scanner features lightweight, carbon fi¬ber alignment rods from Measurement Devices Ltd.'s (MDL) Boretrak MKII. This is a borehole deviation system that stops the two axis measurement head rotating and lowers the system down a borehole. The scanner is deployed by a 110-mm (4.3-in.) borehole, up, down or sideways. The operator can then carry out vertical and horizontal plan section scans at any user specified ARC (an angular increment of 2°) or the distance between two points on the ARC (CHORD) increments. Windows '95 software enables the operator to control the cavity scanner remotely, surveying the area in real time. The system makes data quick and convenient.

Jan 1, 1999

By James R. Arnold

Gold Fields Mining Corporation started processing ore from it's Chimney Creek operation in northern Nevada in November, 1987. Since that time production has averaged about 200,000 ounces per year from the milling and dump leaching circuits combined. The events that led up to commissioning of the plant started in January of 1985 when the Gold Fields exploration team drilled their first hole in an area now known as Discovery Hill. The major orebody was located later in the year in what is currently the major focus of activity, the North Pit. Subsequent drilling since plant startup has found another 14 million tons of ore two miles south of the original orebody. This South Pit orebody is currently being stripped and metallurgical testing is underway. Engineering Engineering on the Chimney Creek plant started in earnest in May, 1986. Davy McKee Corporation from San Ramon, California, did both the engineering and the construction management. Before engineering began the entire senior staff of the Plant Department was hired and assigned to the project. Senior staff included the Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist. Gold Fields does not employ an internal central engineering staff to design their plants. We feel that the people who design the plant should be the ones who will be required to run it once it has been commissioned. This philosophy affords a continuous accountability from start of design through commissioning and operations. All four of the senior staff spent the most critical phase of the engineering stage at Davy McKee, so the plant is truly a reflection of all four disciplines. In bringing on the Maintenance General Foreman from the start, there was a focus placed on ease of maintaining the plant. Another philosophy that was adopted from the start was that the plant be easy to work in and maintain for both operators and maintenance personnel. Numerous design criteria were given to Davy McKee to ensure this. For example, we required that every piece of equipment could be reached easily with a crane or fork lift. All conveyors were required to have walkways on both sides for ease of maintenance and a minimum of 30 inches of clearance under them for ease of cleanup. Visibility for operators was considered very important and the control rooms and general plant layout were designed with this in mind. A list of about 100 general rules were given to Davy McKee during the early layout meetings and these were held as the general design criteria throughout engineering and construction. While engineering was progressing in San Ramon, California, dump leach test work was being done at the site. Three 6 foot diameter fiberglass columns were constructed and filled to a 20 foot height with low grade ore. Leaching characteristics were determined and the dump leach design was based on information collected from these columns. Construction The major contractors on the project were The Industrial Company for the bulk of the piping and structural, EMKO on the electrical/ instrumentation and Hunewill for most of the earth work and leach pad construction. All plant operating and maintenance personnel were employed eight weeks before startup for intensive training. Instructors were Gold Fields personnel, including foremen, and Davy McKee engineers. On occasion, selected suppliers representatives were brought in to discuss major pieces of equipment. Project completion was targeted for January, 1988 at a projected cost of 78.9 million dollars. Actual startup was about 1 1/2 months ahead of schedule and slightly under budget. Full design capacity was achieved the first full month of operation. Crushing Run of mine ore is delivered to the crusher and dumped into a 90 ton hopper which is topped with a 30" X 30" opening stationary grizzly. Ore is fed to a 50" X 60" Telsmith jaw crusher with a 5' wide apron feeder. The closed size setting of the

Jan 1, 1989

By Richard R. Klimpel

INTRODUCTION Froth flotation is the most widely used and economic means of concentrating metal sulfide ores such as those containing copper, lead, zinc, nickel, molybdenum, and pyrite. Also recoverable are other metal species that are often associated with sulfide ores, such as cobalt, platinum, gold, silver, etc. Flotation of metal sulfide ores has been successfully practiced at the industrial scale for more than 50 years. It is the relative technical/economic ease of floating metal sulfide minerals that has been a major driving force behind the rapid expansion in the supply/demand of metals over this 50 year period. It is estimated that 1.6 billion tons of feed sulfide ores are processed each year in the free world based on the froth flotation process, Klimpel (1987) and the manufacture/use of associated chemical reagents is a major portion of the 1.5 billion dollars (U.S.) free world mining chemicals business, Klimpel (1988b). The use of froth flotation appears to be increasing partially due to the ever decreasing feed grades and liberation sizes of the feed minerals. Froth flotation has shown itself to be a flexible process that lends itself well to many solid/solid separations. Thus, with surprisingly little equipment modification, separations can involve very different particle sizes and densities, and relative weight ratios of materials to be separated. In addition, the flotation process is economic, especially when compared to size reduction, its associated precursor process. Flotation also lends itself to continuous operations with a variety of equipment configurations. Such operations exhibit an ability to vary feed rate of solid to the process by as much as 50% without a total collapse of separating efficiency. In addition, the widely used mechanical flotation cell is scalable from 2.8 to 57 or 85 m3 (100 to 2000 or 3000 cu ft) with surprising ease relative to other unit engineering operations over the same relative size increase. The separating medium used is water. Most of the chemicals required -- pH regulators, frothers, collectors, activators, and depressants -- are all relatively inexpensive and common. They are not usually used in large quantities. While there has been a tendency for the quality of metal sulfide ores in any one geographic location to deteriorate over time due to mining intensity, the overall global supply of sulfide minerals is still very ample. Thus, the physical supply side of sulfide ore concentrates has not been under strong technology improvement pressure with increasing metal demands. Part of this is due to the ability of capital to consistently move to new geographic sources having suitable-grade sulfide ores. In these moves, the actual froth flotation technology used (including collector chemicals) has had to change little. As an illustration of this trend, of the approximately 80,000 metric tons of thio collectors used commercially (1980) in the free world, almost 987 were known and manufactured in some form 25 years ago. This is clearly not typical of reagent development in other process-related industries. In addition, the industrial-scale practice of froth flotation applied to sulfide ores has proceeded since the 1920's with often little direct (predictive) scientific explanation due to the extreme complexity of the flotation process. Empirical testing has been a mainstay of industrial flotation reagent development and use. Even today there is often strong disagreement between researchers as to the mechanisms of chemical flotation practices that have been performed successfully on an industrial scale for many years. Thus the industrial process of froth flotation, especially as applied to the recovery of sulfide minerals, poses a dilemma: such a process is widely used, economic, versatile, forgiving, etc. and yet is still not very well understood in a mechanistic sense leading to prediction of results from fundamentals even after 50 years of usage. Lack of such predictive ability has not limited the general over-all industrial use of froth flotation in sulfide mineral recovery. Rather the lack of

Jan 1, 1989

By Louis J. Trostel

Refractories, called the Hidden Industry by The Refractories Institute, provide the temperature and chemically resistant linings for the multitude of vessels that are used today in high temperature processes ranging from giant iron-making blast furnaces to industrial incinerators to petrochemical cracking units to the smallest furnaces used by the jewelry trade. While these refractory linings are absolutely essential to the proper functioning of these processes, the linings are almost always on the inside of the vessels and thus hidden from the eyes of the public. Refractories are defined in Standard C71 of the American Society for Testing and Materials (ASTM), titled "Standard Definitions of Terms Relating to Refractories," as "nonmetallic materials having those chemical and physical properties that make them applicable for structures, or as components of systems, that are exposed to environments above 1 000°F (538°C)." This criterion of ability to withstand exposure to environments above 538OC is the critical distinction separating refractories from other ceramics, fibers, and coatings applicable only at lower temperatures. In addition to resisting temperature, the refractories must with- stand abrasion, chemical and slag attack, resist thermal shock, and carry sustained structural stresses at high operating temperatures for as long as years. The most common methods of classifying refractories are by forming method and composition. First are the traditional formed and fired bricks, batts, and the associated multitude of special shapes. These refractories are usually formed at room temperature, dried, and then fired to high temperatures to develop the final properties desired. Other formed refractories are chemically bonded or resin or tar bonded. To complete the formed refractory matrix, some fired refractory bricks and shapes are subsequently impregnated with resins or tars. Physical dimensions of these formed refractories are classified into two general categories: standard sizes and special shapes. The standard sizes are those in common usage in the United States described in ASTM Standard C909 "Dimensions of a Modular Series of Refractory Brick and Shapes" based on the 38 mm basic module as described in ASTM Standard C861 "Determining Metric Dimensions of Standard Series Refractory Brick and Shapes." The most common size is the "9-inch straight" 228 x 114 x 64 mm (9 x 4 ½ x 2 ½ in.) of the "2 ½ -inch" series or 228 x 114 x 76 mm (9 x 4 ½ x 3 in.) of the "3-inch" series. Bricks of nonstandard sizes are referred to as special shapes. Their use is required in specific furnaces for particular requirements or applications. In addition there is the large and growing group of refractories sold as unshaped mixtures that may be formed into their final shapes at the application site. These unshaped mixtures are sometimes referred to as specialties. The mixtures are prepared to be hydraulically cast, plastic or wet rammed, dry rammed, or dry vibrated into their final shapes. Some are designed to be projected through nozzles (or gunned) onto their forms. The refractory mortars used to lay up the shaped refractories fall into this group referred to as specialties. Prepared refractory grains also are sold. These are usually calcined or fused materials that are subsequently crushed and sized to a variety of specific carefully controlled grain sizes or size distributions. These grains include calcined fire clay, dead burned magnesite, periclase (fused magnesia), dead burned dolomite, kyanite, tabular and fused aluminas, fused zirconia, fused mullite, and calcined alumina-magnesia spinel. The primary composition groups of refractories are the 1) alumino-silicate refractories made largely from fire clay and high- alumina materials, 2) silica refractories made from quartzites and ganisters, and 3) basic refractories from magnesia (or magnesite) and chromia (also referred to as chrome ore), alone and in combination, and dolomite. Special refractories include carbon, zirconia, zircon, high purity alumina, mullite, silicon carbide, silicon nitride, and some borides. In addition there are groups of special insulating refractories.

Jan 1, 1994

By James M. Barker, Richard O. Y. Breese

The recognition of perlite as a distinct volcanic glass, and of the arcuate fractures that often characterize these natural glasses, dates back at least to the late 19th century and perhaps into the 18th century (Howell, 1974). Some authors suggest that recognition of perlite dates considerably further into antiquity, perhaps to the 3rd century BC (Caley and Richards, 1956). Most recent definitions identify perlite by the presence of vit¬reous, pearly luster and by the presence of characteristic concentric or arcuate perlitic fractures. As discussed more fully below, arcuate fractures in perlite need not be megascopic. The textures of perlite which commonly occur in deposits range from dense to highly vesiculated to pumice-like. Thus, the classical definition of perlite is descriptively restrictive compared to the natural range of textures and neglects to address the issues of genetic origin and geologic occurrence. In spite of the petrological usage of the term perlite, the in¬troduction of expanded perlite aggregate into industrial markets caused the term to be applied as well to the lightweight, cellular aggregate that is produced through the rapid thermal expansion of milled perlite ore. Perlite is distinguished petrologically from other natural glasses by a silicic or rhyolitic composition by the presence of 2 to 5 wt% total chemical water held within the glass structure and often by the presence of pearly luster and onionskin-like perlitic fractures. Worldwide, occurrences are associated with Tertiary through early to middle Quaternary volcanics and with the glassy portions of silicic domes and flows, with vitric tephra, with the glassy chill margins of dikes and sills, and with the glassy portions of welded ash flow tuffs. Less commonly, perlite is reported in association with the glassy portions of volcanic plugs and lacco¬liths. The perlite carapaces that partially or fully comprise extru¬sive domes and flows are often thick and areally extensive, and thus provide commercially attractive targets for low cost open pit mining. Perlite tuffs and tephras are also important commercial sources. Mining costs are minimized by open pit quarrying and by either bulldozer ripping or blasting. Crushing and sizing facilities are generally close to the pits. Due to the low weight and large volume of expanded perlite, unexpanded perlite that has been crushed and sized is usually shipped nationwide directly to local markets where it is expanded and processed for distribution to end users. The vast majority of international trade is in unexpanded perlite. In the United States, the evolution of the expanded perlite industry can be traced to the late 1930s and early 1940s when expanded perlite aggregate was introduced into gypsum plaster and concrete markets as a substitute for vermiculite (Schundler, Jr., oral communication, 1991) and was promoted for use in a much broader spectrum of products (Howell, 1974). By the early 1950s, end user markets were well established and small expanders were numerous east of the Mississippi River. Today, mining and expan¬sion is undertaken throughout the world. Worldwide production of perlite is dominated by the former Soviet republics, the United States, and Greece. Total worldwide production of processed perlite during 1990 is estimated at roughly 1 778 000 tons (Davis, 1991). The physical properties of expanded perlite that give this prod¬uct unique commercial value include: low bulk density, chemical inertness, high insulating ability, nonflammability and fire resis¬tance, and its ability to retain water. These and other physical properties give expanded perlite value in such uses as: lightweight acoustical and thermal insulating aggregate; component in plaster, insulating cements and lightweight concrete; loose fill insulation; as aggregate in horticultural applications; and in milled form as filter aid. In the United States and Europe, the largest end use is within building construction products.

Jan 1, 1994

By Richard L. Bullock

INTRODUCTION TO COAL MINE PRACTICES To illustrate the trends and the proportions of the various methods used to mine coal underground, Fig. 1 is useful. However, of all the underground coal mining practices in the US, room-and-pillar mining still accounts for 96.5%. This percentage is divided as follows: conventional, machine cut-33.0%; conventional, hand cut -1.7% : and continuous mining-61 .8%. As was stated in Sec. 1, Chap. 7, the term "development" means different things to a coal miner and a metal miner. Though some of the ideas and concepts presented in the first two chapters of this section are valid for all types of mines, they apply particularly to mines using noncoal methods. In coal mines (and many other mines using coal mining systems) development refers to that phase of the mining cycle which advances several entries through the virgin seam, forming pillars by cross- cutting between the parallel entries. Because in the past, coal seams (as well as seams of potash, trona, oil shales, gilsonite, salt, and some uranium) have been much more uniform in their thickness, continuity, and quality, and because they often cover large areas, it has been possible to develop the room-and-pillar mine completely from a preplanned, systematic approach with a minimum of surface drilling information. As the better coal seams of the midwest and east are being depleted (or eliminated because of high sulfur), seams of poor quality and continuity will have to be mined. Also, seams that are not nearly so well known at greater depth will have to be taken as well as some of the steeply pitching and folded seams of the west. The American coal miner is only now feeling the need to do a much more complete job of prospecting the geology of the coal seams than he has ever done before. This is the only way that the risk of unknown rock conditions can be minimized. In many ways, surprises in the rock or coal seams are much more of a disaster to the coal miner developing a room-and- pillar mine than to his metal mining counterpart. Be- cause of the multiple entries, the complex set of the Federal Coal Mine Health and Safety regulations, and the totally dedicated design of coal mining equipment, coal mining systems do not have the flexibility to easily overcome changes in the seams mined or the rocks around them. Thus, if a wide parting of rock shows up unexpectedly in the middle of a coal seam entry being milled away by a continuous miner, it will either have to slowly "chew" through it, convert to a conventional mining system, or stop that particular entry. Likewise, if suddenly the coal disappears completely and in its place are the remains of an old creek bed, the "want" or "wash out" as it's called will usually have to be mined right along with the coal, by the same methods and with the same soft rock equipment, or the entry will have to be abandoned. For reasons such as these, coal seams today should be drilled from the surface just as thoroughly as the ore bodies in metal mines. Having once determined a more precise location of where the coal seams are located and which are suitable for mining, the engineer must still take a systematic approach to planning the mine as completely as possible. The initial series of entries must be planned to be driven either from the side hill portal or from the bottom of the shaft or slope. In either case, they should usually intersect the area close to the midpoint of the mass to be mined.

Jan 1, 1982

By Yuting Xue, Brijes Mishra

The relaxation test on sandstone specimens showed typical behavior in pre-failure region and stepwise behavior in the postfailure region. Overall, more significant stress relaxation occurred within the failed specimens than the intact ones. Numerical simulations were conducted with pre-defined failure plane and with Voronoi tessellation to visualize the relaxation behavior. The model with pre-defined failure plane showed the key role of failure plane and asperity in simulating the step-wise post-failure relaxation behavior. The inhomogeneous stress distribution within failed specimen and the viscous deformation of intact rocks induced high stress concentration at the asperities. The observed step-wise relaxation initiated from the failure of asperity. Furthermore, the simulations with Voronoi tessellation showed time-dependent fracture development during relaxation in post-failure region. The presence of fractures completely changed the stress distribution. Stress concentration occurred at the front area of fractures and at the interacting areas between fractures. During relaxation, fractures still developed with time and the sudden significant increase in the “damage” coincided with the acceleration of stress relaxation leading to step-wise relaxation. Finally, the results showed the possibility of using residual strength as the long-term strength of failed rock.

Jun 25, 2020

By E. Galán

The largest Spanish kaolin deposits formed on granites occur in Galicia (NW Spain) due to a widespread kaolinization of Variscan granitoids. This paper provides insight of the origin of this kaolinization process on the basis of a mineralogical and geochemical study carried out in two kaolin deposits, presently mined to be used in the paper industry. Mineral composition of the deeply kaolinized rocks is dominated by well-ordered kaolinite with minor quartz, K-feldspar, mica (illite/muscovite), locally gibbsite, and traces of resistant accessory phases (rutile, ilmenite, zircon and monazite). The content and structural order of the kaolinite decrease with depth. Kaolinite formed by extensive replacement of feldspars (plagioclase). Differential mobilization of major and trace elements during kaolinization has been inferred by using a mass-balance approach. Significant amounts of Na, Ca and minor Mn, Sr, P and U were removed from incipiently leachable minerals (plagioclase and apatite), and K, Rb, Cs and Ba were released from K-feldspar and mica. On the contrary, Ti, Zr, Th, Hf and the REE are retained in the resistant trace minerals that are concentrated in the fine fractions of the kaolins.

Jan 1, 2010