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By J. B. Gwiazda

S.S. Peng The highlights of this paper are: The shield supports with lemniscate links will exert a tangential force on the roof strata, causing unnecessary and harmful roof movements, thus damaging the roof structure. The principal source of the tangential force is the horizontal movement of the point O, the pivotal joint between the canopy and the caving shield. When the support height is changed during setting, the point O will move along a curve upward and rearward toward the gob. As a result, it produces a vertical movement, causing the roof to move up a distance of t d, and a horizontal movement toward the gob, creating vertical fractures in the roof. The major purpose for employing the lemniscate- link system is to protect the hydraulic legs against the action of the bending moments induced by the tangential force that causes deformation or destruction of the legs. Based on these two points, the author suggested that, in order to eliminate the harmful effects of the lemniscate-link system and strengthen its advantageous features, a shield support with a new design of that mechanism has been developed. With the new design, the conventional rigid lemniscate links are replaced by hydraulic rams. When the pressure inside the rams reaches a certain level, the ram will yield.

Jan 1, 1987

By A. Iordanidis

Ieropigi Fe-Ni deposit is located in northern Greece, in Kastoria?s prefecture near Albania?s borders. A laterite profile exposed by opencast mining and placed near the basement of the mine was chosen for sampling. Ten samples, representing ?pure? Fe-Ni ore, ophiolite-bedrock, ?red? ophiolite and ?enriched? ophiolite were collected, based on macroscopic observation. The samples were analysed by X-Ray powder Diffraction (XRD) for the determination of the mineral phases. Iron and Nickel concentrations of the same samples were also determined by X-Ray Fluorescence (XRF) methodology. The Fe-Ni ore samples consists predomi-nantly of hematite, goethite, chrysotile (serpentine), saponite and quartz. ?Red? ophiolite samples consists mainly of chrysotile, hematite, quartz, calcite and small amounts of paragonite and goethite. Bedrock and ?enriched ophiolite samples are comprised of chrysotile, hematite, calcite, quartz, nontronite, paragonite and sepiolite. The Fe content ranges between 13% and 49% in Fe-Ni ore samples and between 7% and 10% in ophiolite samples. The Ni content ranges between 1% and 3% in Fe-Ni ore samples and between 0.3% and 4.5% in ophiolite samples (the higher content is observed in the so-called ?enriched? ophiolite). A correlation between Fe-Ni concentrations and mineral assemblages is being attempted in order to evaluate the nickel-rich mineral phases. It is shown that the nickeliferous minerals are serpentine, hematite, goethite and saponite.

Jan 1, 2006

By J. Moreno, P. Sinton, R. Vogwill

Groundwater models are often used to estimate and optimize dewatering volumes, locations, and rates, as well as water supply production for mines. Projections of future dewatering rates are highly dependent on the conceptual model used as a basis for the numerical model. This paper presents an audit of modeled dewatering rates versus actual data and an examination of the reasons for discrepancies. The setting is a gold mine, in a collapsed volcanic crater breached by the ocean. Mining will occur to about 200m below sea level within a remnant geothermal system. A detailed hydrogeologic investigation was conducted to support dewatering and geothermal depressurization design for future mining operations. These studies included development of a three-dimensional model to simulate heat and density-coupled groundwater flow, using software developed by URS/Dames & Moore. In this paper, the groundwater modeling results prepared 5 years previously by others was audited using operational dewatering data. This model was then extended using the new mine schedule and actual locations for dewatering wells to predict future dewatering requirements. In addition, a comparison between “hot” and “cold” model results was used to assess the need to use a heat-coupled versus a standard groundwater model for dewatering predictions.

Jan 1, 2003

By J. E. Murphy, Y. Topkaya, R. E. Johnson, B. R. Palmer, D. K. Dysinger, R. B. Bhappu, A. E. Clark

The U.S. Bureau of Mines utilized a central composite experimental design to investigate leaching kinetics of pure chalcopyrite in the sulfuric acid-oxygen system. The six factors studied were oxygen pressures of 150, 300 and 450 PSI, leach temperatures of 35O, 65O, and 95O C, oxygen-to-chalcopyrite mass ratios of 0.5, 1.0 and 1.5, leach times of 8, 16 and 24 hours, initial acid strengths of 0.6, 1.2 and 1.8 g H,SO,/g chalcopyrite and pyrite to chalcopyrite mass ratios of 0, 1, and 2. Chemical analysis and oxygen consumption measurements indicatedthat leaching occurs through three consecutive reactions. Temperature was the primary determinant of leach rate and solution acidity was the primary determinant of oxygen consumption, sulfur oxidation, and activation energy.

Jan 1, 1993

By Donald G. Foot

The Bureau of Mines, U.S. Department of the Interior, devised a direct flotation method for treating low-grade carnallice (KC1.MgC12. 6H20) and sylvinite (a mixture of KC1 and NaC1) ores characterized by a high content of water-insoluble impurities. The procedure consists of (1) flocculation-depression of the insoluble slimes, (2) decomposition leach (carnallite only), (3) potash rougher, cleaner and recleaner flotation, and (4) final product leaching. The technique eliminates the necessity of removing the water-insoluble slimes by mechanical or flotation desliming. Laboratory results demonstrated that the direct flotation procedure, when compared with conventional techniques, yielded equivalent or improved flotation results with equal or less reagent consumptions. The direct flotation method recovered 81.6% of the solid-phase KC1 from carnallite ore and 79.7% of the KC1 from sylvinite ore in final products containing 60.5% and 60.9% K20, respectively. A flowsheet incorporating the described flotation procedure is presented.

Jan 1, 1981

By Frank Florez

INTRODUCTION The Magma mine, located at Superior, AZ, some 100 km (60 miles) east of Phoenix, was discovered in 1875, and was originally mined for silver. The property was acquired by W. B. Thompson who formed the Magma Copper Co. in 1910. Prior to 1965, square-set mining and cut-and-fill methods were used to mine steeply dipping veins at Magma. Ore was moved to the chute by gravity in the square-set stopes and by slusher in the cut-and-fill stopes. The stopes were gobbed with waste from the surface and development headings in the mine. The original square-set method of mining was modi¬fied in order to extract replacement ore bodies in limestones that dipped at 0.52 rad (30°) and varied in thick¬ness from 4.6 to 38 m (15 to 125 ft). Mining of these ore bodies was done by starting at the upper lateral ex¬tremities and retreating to the access openings. The timber square sets and cribbing were used as temporary support, the ground being allowed to cave as the mining retreated. It was found, however, that mining in such a manner was very inefficient; safety hazards increased be¬cause of the blocky and fractured characteristics of the ore and the weakness of the hanging wall. GENERAL DESCRIPTION The undercut-and-fill method of mining is now the primary stoping method used at Magma, and no pillars are left for later extraction. Four ore bodies are pres¬ently being mined. These ore bodies (see Fig. 1) dip conformably at 0.52 rad (30°) to the east with the host limestone and shale sediments, their average strike length is about 122 m (400 ft), and their thickness varies from 9 to 38 m (30 to 125 ft). The ore bodies are separated by varying thicknesses of barren rock. Three of them come together at various elevations primarily in the area of the main feeder fault, but there is an apparent horizontal displacement of about 23 m (75 ft) on the main feeder fault which complicates stope layout. The method as originally applied has been modified in order to mine more efficiently ore bodies that are 9 m (30 ft) or more in thickness. The modification was made to mine all the ore at a given elevation before dropping down for the cut below. The advantage of this is that mining is done on a horizontal plane and there¬fore muck removal can be mechanized. Also multiple faces are made available, thereby improving overall min¬ing efficiency. Another major change made was to carry a 6.7-m (22-ft) face when mining the top sill and a 6.4-m (21-ft) face when mining under fill rather than the former standard 3.35-m (11-ft) high face. This change lengthens the mining cycle and cuts mat prepa¬ration costs in half. Stopes are laid out (see Fig. 2) so that maximum mining panels are available and yet one stope crew will not tie up too much of the ore reserve. One stope might consist of as many as 20 panels, and these panels are divided evenly on opposite sides of the extraction cross¬cut. The stope layout is planned so that the maximum tramming distance to the orepass will not exceed 30.5 m (100 ft) on the crosscut and 38 m (125 ft) in the panel. Stope panel distances are sometimes exceeded when min¬ing to the extremity of the ore body. Tramming distance limits are determined by the efficiency of the com

Jan 1, 1982

By Joseph W. Leonard

Much progress has been made in the application of geostatistics to hard rock ore deposits.1,2,3 4,8 Progress in the application of geostatistics to coal deposits has developed over a period of time out of hard rock technology.5,7 Exploratory drilling in many hard rock deposits is frequently significantly more productive than exploratory drilling in coal deposits. For example, hundreds of feet of drilling and/or many drill setups are frequently necessary to obtain a few samples from a typical 4 to 6 foot thick coal seam. With hard rock deposits, which are frequently thick or volumetric, less drilling dead work is involved in sampling the deposit. To produce a variogram, which is the basic element needed for geostatistical analysis, a few well positioned holes drilled into a hard rock deposit where the core(s) can be divided into many sections can produce a detailed variogram. To produce a similar type variogram in coal, it is necessary to drill many dozens of holes with much of the core length consisting of overburden data which is frequently of only secondary importance. To apply geostatistics to coal, not only are many separate sample sites needed, but frequently a problem exists in that the interval distance between samples cannot be controlled with the ease that hard rock drilling intervals can be controlled where it is only necessary to cut core segments according to some preassigned interval of distance. Because the uses of existing coal deposit samples are frequently handicapped by irregular sample distances and lack of sampling density along certain directions needed for the development of variograms, the application of geostatistics to coal deposits frequently does not proceed with the same efficiency and rapidity as when these techniques are applied to ore deposits. Grid-fitting was developed by the writers, as a logical and evolutionary improvement over the long and successful application of geostatistics to coal seams.6,9 Grid-fitting combines the trending of variance as well as other aspects of probability theory to optimize informational yields from the commonly less productive data derived from coal exploratory programs. Grid-fitting is basically a technique where a hypothesis involving numerous grid patterns, where each pattern is of a given geometry, size and direction is repeatedly matched against actual sample data to determine which match results in the closest fit of the hypothesis to the data. To determine the closest fit, a special index known as a selectivity index was developed. The selectivity index is basically the product of the mean and variance of the variances contained within each of the grids. hence, if SI is the selectivity index then [SI = Xs E (S i - Xs) /N] where [X]s is the mean of individual grid variances Si2 for any pattern of given geometry, direction and size.

Jan 1, 1984

By Luther W. Hyde

Ground water, one of Florida's most valuable resources, is also one of its most abundant. The State is fortunate to have several prolific ground- water reservoirs that yield tremendous quantities of water to wells and to some of the world's largest springs. These ground-water reservoirs, or aquifers are of two types artesian and nonarteaian. Ad artesian aquifer is one that contains water under sufficient pressure to rise above the top of the oontaining formation. A nonartesian aquifer is one that contains water which is not confined, and the upper water surface (water table) is free to rise and fall. The principal source of ground water in Florida is the artesian Floridan aquifer which supplies most of the water users in the State. Nonartesian sources include the Biscayne aquifer of southeastern Florida, the sand-and-gravel aquifer of extreme western Florida, and a shallow aquifer which lies at depths of less than 100 feet over much of the State. The accompanying map shows the approximate areas where these aquifers are the main source of supply; however, the boundaries do not represent the areal extents of the aquifers. For example, the Floridan aquifer extends beyond the southern boundary shown on the map, but it is not of great economic significance there because the water is highly mineralized. Aquifers other than the four described occur in the State, but they are, for the most part, of minor significance. One of these is an artesian aquifer which is formed by permeable zones that occur locally within the confining beds above the Floridan aquifer.

Jan 1, 1966

By R. R. Fernández

Newmont Mining Corporation has been processing refractory gold ores by roasting since January 1995. The ores are refractory due to the association of the gold with iron sulfides, namely pyrite and marcasite, and the presence of carbonaceous components and is oxidized in an oxygen-enriched gaseous atmosphere. The structural characteristics of the calcine have a strong influence on the eventual recovery of gold by cyanidation. These characteristics are strongly dependent upon the roasting conditions, mainly the roasting temperature. Originally, the sulfides and carbonaceous matter were oxidized at a temperature that typically ranged between 535 and 560 °C. Controlling the roasting temperature within the lower and narrower range of 510 to 515 °C produced a calcine with a larger surface area(e.g. higher porosity), which resulted in a gold recovery increase of3.3 percent in 2001 equivalent to 5.1 million dollars of additional gross revenue. The new strategies of operating at temperatures never tested before are discussed. It constituted a great challenge to the control room operators for the lower operating temperature falls much closer to the ignition temperature of the sulfide ore and smaller and more frequent adjustments in operating parameters were needed to improve and maintain high gold recoveries. Based on the encouraging results that were achieved in 2001, a decision was made to install a “smart” process control system because the roaster temperature control was performed by manual inputs from the control room operators. The development and implementation of the control system to reduce or eliminate the temperature variability between the different control room operators commenced in June of 2002 and was fully operational by the end of October. Lower roaster temperatures were achieved (500 to 505 °C), which resulted in further increase of gold recovery by about two percent. The overall recovery increase equated to 8.0 million dollars of additional revenue for the first 10 months of 2002.

Jan 1, 2003

By Cal D. Iles

Decreasing ore grades and increasing labor and material costs have seriously eroded the economic vitality of many mining operations. Coupled with global economic stagnation and recession in the early 1980's, these conditions have forced reevaluation of operating strategies and methodologies with the objective of remaining competitive in a world marketplace. A concept offering large potential benefits is hardrock portable crushing and conveying. Implementation of this system at Duval Corporation's Sierrita open pit copper-molybdenum mine (Figure 1) has significantly reduced mining costs, becoming a major part of Sierrita's cost reduction strategy. Located 48 km south of Tucson, Arizona, the 209 000 t/d property has one portable 60x89 gyratory ore-crusher in operation with tentative plans for two additional units.

Jan 1, 1985

By W. A. Haley

Wheel loaders became important bulk materials handling tools in the early fifties. As they grew in stature, productivity, reliability, and acceptance, the demand for larger, more rugged machines evolved. In the early sixties, the introduction of a 6-yard model by a major equipment manufacturer led quickly to widespread experimenting and rapid worldwide acceptance in loading blasted rock in stone quarries and smaller metal mines. As the wheel loaders grew in popularity for shot rock loading, larger models, 12- to 15-yard size were developed and saw widespread use by 1970. In the past 20 years, at most quarries and small to medium size mines, most cable operated power shovels have been replaced by large wheel loaders and more recently, hydraulic excavators as the primary loading tool. Wheel loaders larger than the 15-yard class are not used widely. Until about 1970, very few overburden haulage trucks were used in domestic coal mines. About that time, many states started enforcing stringent mine relcamation regulations and in 1977, even more stringent federal legislation was passed. These controls brought about drastic changes in coal surface mining methods in the mountainous areas of Appalachia, in smaller midwestern mines and in the large mines of the western states. Basically, simple material casting methods were replaced at most eastern mines by "haul-back" and "mountain top mining" methods using controlled placement of overburden as it is removed from the coal. These methods dictated much more extensive use of loaders/ shovels and trucks to the point that most coal mining in those areas is now by loader/shovel trucks methods. And those areas have become the home of the country's largest population of large wheel loaders and trucks in the 35- to 85-ton size class with a trend to even larger trucks and shovels. To maintain equal production or provide for growth, the new methods have caused a sharp increase in machine intensity to the extent that at some mines, up to three times as many machines were required to mine a million tons of coal in the late seventies as compared to the machine needs in the sixties. Mining costs and coal prices also took a big jump in the mid-seventies, paralleling the drastic price increases in oil and other forms of energy.

Jan 1, 1984

By James J. Scott

Three dimensional photoelastic techniques have been developed which provide information on the state of stress at any point in a model investigated. This investigation utilizes these techniques to determine the stress distributions in the pillars and the roof of a room and pillar mine model. Castolite plastic was used for the construction of the model in which a stress pattern was developed by the stress freezing process. Analysis of the stress pattern was accomplished by use of data obtained by observing points in slices from the model with a polariscope and solving data obtained by' the shear-difference method. Slices through the centerline, edge of pillar and centerline of span were investigated. Results of this work provide information on the overall stress distribution on points of stress concentration in compression, shear and tension, show relative magnitude of stress throughout the model and indicate directions of the most important stress values. A generalized roof bolt pattern is developed for room and pillar mines based on the stress distribution. A digital computer was used in the reduction of data and proved to be extremely useful for investigation of this type.

Jan 1, 1963

By C. A. Rowland

Introduction Rod mills are horizontal, cylindrical, rotating, tumbling mills in which steel rods, that are usually about 6 in. shorter than the length of the grinding chamber, are used for grinding media. Fig. 33 shows a cutaway view of a rod mill. Rod mills were first used around 1910 because of their ability to grind in open circuit, consistently produce a uniformly sized prod¬uct, give greater continuous availability, and give lower maintenance costs. They have basically replaced roll crushers and coarse grinding ball mills in comminution flowsheets. The present 20-ft limit on rod length, based on rod manufacturing methods, and the pulp flow char¬acteristics, that influence mill capacity, have established the maximum rod mill sizes' of 15 x 21 Vs ft using 20-ft rods and drawing from 2,200 to 2,300 hp. Rod Mill Uses. The principal uses for rod mills are: preparing feed for ball or pebble mills, grinding gravel or aggregates to the size analysis required for making concrete sand, preparing coarse agitator feed for leaching operations, and for primary grinding circuits, particularly in industrial mineral applications. Wet grinding is the most frequently used grinding process for rod milling. There are a few special cases where dry rod milling is used. Preparing feed for ball or pebble mills is the most common use for rod mills. In ferrous and nonferrous ore concentrators, wet open-circuit grinding is used to prepare ball-mill feed. Figs. 34 thru 37 show the four basic grinding circuits used in these concentrators. In rod mill-ball mill grinding circuits, where the rod-mill discharge contains 25% or more of the grinding circuit product-size material, the rod-mill discharge can be fed to the ball- or pebble-mill classifier as shown in Fig. 34. This is the most frequently used rod mill-ball mill circuit in the nonferrous ore concentrators. In a few concentrators such as Anaconda's Weed concentrator,55 the rod mill discharges to one classification stage which makes the same or a slightly coarser particle-size cut as the classifiers used to close the ball- or pebble¬mill circuits, as shown in Fig. 35. In concentrators where rod mill-ball mill circuits are used to grind low grade magnetic iron ores such as Reserve Mining Co., Silver Bay, Minn.; Erie Mining Co., Hoyt Lakes, Minn.; Eveleth Taconite Co., Forbes, Minn.; US Steel, Mt. Iron, Minn. and Atlantic City, Wyo.; Inland Steel-Jackson County Iron Co., Black River Falls, Wis.; and The Cleveland-Cliffs Iron Co., Adams mine, Kirkland Lake, Ont., the rod mills discharge to rougher magnetic separators (cobbers), the concentrates from which are fed to a ball milling stage as shown in Fig. 36. The same basic circuit is also used with gravity concentra¬tion methods such as spirals (Hanna Mining Co., Groveland, Mich.), tables (St. Joe Minerals, Bonne Terre, Mo.), and jigs. When the rod-mill discharge contains less than 25% of the grind¬ing circuit product size, this indicates an ore that does not readily produce the fines needed to help carry the coarse feed into the ball charge. When this occurs, the fines in the rod-mill discharge are needed in the ball mill or pebble mill so the rod-mill discharge flows direct to the second-stage mills as shown in Fig. 37. Rod mills are used to make feed for open-circuit ball mills such as those used in wet grinding cement raw material (American Cement, Brennan Ave¬nue plant, Detroit, Mich., and Port Huron, Mich., plant, and at the South Dakota State Cement plant, Rapid City, S. Dak.). Grinding gravel or aggregates to the size analysis required for making concrete sand is used in areas where natural sand of the proper particle size gradation required to make concrete is not availa¬ble. Three-quarter inch crushed aggregate is ground to 4 mesh using center peripheral-discharge rod mills (Fig. 38) operating in open circuit. US Corps of Engineers specifications state that the sand product be 93-100% passing 4 mesh and contain no more than 3-7% passing 200 mesh. It may be necessary to classify the rod-mill discharge to remove any excess -200 mesh sand produced in the rod mill. Particu¬larly when processing aggregate with high abrasion indices, center peripheral-discharge rod mills are used instead of hammer mills, fine bowl crushers, and crushing rolls for the production of specification sand. If the -4 mesh size fraction of the rod-mill feed is removed

Jan 1, 1985

By Michael R. Grubbs

In most resource recovery plants the incoming solid waste material is first shredded and then air classified into a light and a heavy fraction. The light fraction which contains mostly paper, cardboard, etc., can be used either as a fuel for power generation or in the production of new paper products. The heavy fraction which remains consists mostly of glass, ceramics, wood, ferrous metals and non-ferrous metals. The ferrous metals may be removed from the heavy fraction by conventional magnetic separators. After primary shredding, most of the glass and ceramic is fine in size and may be screened away leaving a residue of wood, cardboard, and non-ferrous metals. This residual heavy fraction is the input material for the separator described in the present paper. A new metal separator was recently reported by t. Schlömann1. The author described the theoretical relationships and presented results of tests on samples of the residual fraction of municipal solid waste. This paper describes the operation of the metal separator and develops the basic separation relationships controlling the process.

Jan 1, 1976

By R. Kramer, A. Habibi

"Diesel Particulate Matter (DPM) measurements were taken at Tronox’s Westvaco Trona mine in Wyoming. Both shift-average and real-time monitoring using Elf pumps and Airtec analyzers, were used. The mine’s underground transportation fleet consists of more than 250 diesel-powered vehicles. A vehicle monitoring log tracked fleet usage during the test; logged data consisted of vehicle type, direction of travel, and time. The log was used in analysis to determine the DPM spread rate, of normally operating equipment, in the different entries with respect to vehicle location. Airflow velocity measurements at the stations were taken during the survey. The survey results also identified vehicles with higher emission rates. An ECOM gas and a MAHA DPM analyzer, both located in the mine’s underground emission shop, were used to obtain various vehicle emission rates. The correlation of the workshop and field measurements was investigated with regard to the machine activity, fuel usage and atmospheric conditions. The results were used to identify and execute a ventilation project to reroute 45% of diesel shop air back to intake entries, by interlocking an automated door and a regulator. The result is reduced DPM downstream. INTRODUCTION DPM has been classified as a potential occupational carcinogen by the National Institute for Occupational Safety and Health (NIOSH).The Mine Safety and Health Administration (MSHA) rule governing DPM exposures in metal and nonmetal mines cites total carbon (TC), a summation of elemental carbon (EC) and organic carbon (OC) obtained by NIOSH Method 5040, as a surrogate for determining DPM exposures (Noll, 2006). MSHA personnel use the 1973 sampling criteria outlined by the American Conference of Government Industrial Hygienists (ACGIH) for collecting compliance-based respirable and inhalation samples in metal/nonmetal mines. Instrumentation developments are allowing improved real time monitoring of ventilation parameters. DPM is within the aerodynamic particle size range of the samplers used for respirable and inhalable dust measurements. If a sampler used by MSHA to collect inhalable or respirable dust samples in underground metal/nonmetal mines was used to collect DPM, both the DPM and mine ore dust would accumulate on the sample filter (Noll, 2006). To comply with the requirement and lower the DPM exposure, mines are implementing a variety of control technologies such as emission based maintenance programs where adjustments or repairs are performed on the vehicles (McGinn, 2000)."

Jan 1, 2016

By Robert K. Jones

Filter media can be defined as that material which effects the removal of finely divided solids from a fluid. Thus the removal of dust from air or turbidity from water, for example, can be the result of passing the fluid through a properly supported filter media. Since filtration in one form or another has been practiced since pre¬historic times filter media have been made from an extremely wide variety of materials. Vegetable and synthetic fibers (used loose, or formed into papers, felts or cloth) sand, ground coal, diatomite, perlite, wire cloth, porous metal, and bonded aluminum oxide are among the more important materials used today.

Jan 1, 1961

By B. Mishra, K. Savage, T. Evans, A. Noble

"Recent government initiatives and commercial activities have targeted asteroids for in-situ characterization, manipulation, and possible resource extraction. Though emergent technologies may be effective for such undertakings, uncertainties remain concerning the effects a space environment may have on the mechanical behavior of rocks. Various non-destructive rock characterization techniques can empirically predict in-situ rock strength on Earth; however, these techniques must be adapted to account for unique space conditions. In order to mimic one element of a space environment, various rock samples were evaluated after prolonged exposure to low temperatures ranging from -75°C to 0°C. Tests included uniaxial compressive strength, Schmidt hammer, and ultrasonic pulse velocity, while rock types included shale, limestone, and chalk. This data was used to generate a temperature correction for the non-destructive methods and evaluate the accuracy of these methods in predicting strength values. Overall, these results provide insight on the mechanical behavior of rocks at low temperatures and may be used to design robust systems needed for asteroid rock characterization INTRODUCTION NASA and various commercial entities have recently targeted near-Earth asteroids for exploration and scientific investigation to better understand their in-situ properties, mechanical characteristics, and formative processes. In particular, NASA’s Asteroid Initiative encompasses several studies and mission plans that will enable future investigations while vetting technologies that can be applied to deep space exploration. One component of the Asteroid Initiative is the Asteroid Redirect Mission (ARM). This multi-stage space mission proposes to rendezvous a robotic spacecraft with a large near-Earth asteroid. Upon approach and after landing, the spacecraft will prospect the surface and identify a multi-ton boulder (up to 5 m in its largest dimension) suitable for capture. The robotic spacecraft will engage the boulder, remove it from the parent asteroid, and return the sample to a controlled lunar orbit. Once the asteroid mass is in stable orbit around the moon, astronauts aboard the Orion spacecraft will engage with the retrieved asteroid for manned exploration that will include sample collection and geologic prospecting. Launch plans for the ARM robotic craft are expected as early as 2020, and manned explorations of the asteroid mass in lunar orbit are expected to follow sometime in the mid-2020s (NASA, 2015). Overall this mission will demonstrate NASA’s ability to intersect, manipulate, and redirect an extraterrestrial body, while supplying scientific information that is relevant for planetary defense and in-situ resource utilization (ISRU) applications."

Jan 1, 2016

General. Data for the operating information given in the 15 tables in this chapter was obtained from voluminous questionnaires sent to over 100 operating companies in 1972. There has been a considera¬ble change since then in supply and power costs and many companies have had expansions or revisions since that time, so it must be realized that information is valid for the year 1972. This is true except for the three new properties of Sacaton, Metcalf, and Pinto Valley. These three operations started up since 1972 and all information for them was obtained in 1975. The questionnaires were sent to all major copper properties throughout the world as well as to many smaller units and to some mines that had complex or unusual conditions. Properties for which no information is given did not submit completed questionnaires. Several operations were unable to supply information requested be¬cause of company policy prohibiting it. However, data is presented for 75 different operating properties. The reference number of the left-hand column is the same for all 15 tables of data. If the company had several operating properties as of 1972, then subsequent properties use the same number but a different letter. For example, Anaconda's property at Butte, Mont., is IA, whereas Anaconda's (now Anamax) Twin Buttes property in Arizona is IB. Table I gives general data for the properties including exact loca¬tion, milling rate, pertinent assay values and recoveries, and a brief listing of ore and gangue minerals. It was not possible to obtain any more information from the properties in Chile except that given in Table I. Crushing. Crushing plant operating data is given in Table 2. This table gives data on the crushers, feeders, and screens used at each property. The data obtained on feeders was confusing, but, rather than completely eliminate this information, it was decided to report it as given. The confusion lies only in the description of where the feeder is located. Where a primary feeder is listed it was not possible to determine if this feeder was on primary feed or primary discharge. For secondary feeders it is known that the feeder listed usually is feeding the secondary crushers, but at some properties it is handling the secondary crusher discharge. The same reasoning holds true for tertiary feeders-they handle either tertiary feed or discharge. This table contains a voluminous amount of data. A column giving tons per day per square foot of screen area would have been very informative, but some properties included circulating load in the feed rate and others listed new feed only, so calculations were meaningless. Autogenous Mills. The use of autogenous or semiautogenous mills for any property must be given serious consideration in the future. Table 3 lists data for four properties that are fully autogenous and four more semiautogenous mills that use balls to supplement the chunks of ore. This method of grinding is especially applicable when the ore is of such a nature that crushing problems would be severe, such as when ore is wet or sticky or contains a lot of fines. Paradoxically, it should also be considered when ore is tough and hard and blocky because then the competency of the ore means it can serve well as the grinding media. Where the ore varies greatly in hardness, as at Cyprus Pima and Lornex, semiautogenous grinding has been especially beneficial with capital costs and operating costs both 10-20% lower than could be obtained with conventional crushing and grinding. Rod Mills. For a great many years the standard milling operation used two- or three-stage crushing followed by open circuit rod mills and then ball mills in closed circuit with classifiers of some kind. A surprising number of properties crush to the fine size of about 3/8 in. and then use ball mills only, in closed circuit, with no rod mills. It is difficult to say whether rod mill-ball mill circuits or ball mill only is the best. One can be certain of two things: (1) the argument will go on for a long time and (2) for some mills one method obviously was superior over the other. As ball mills of 14, 15 ft, and larger came into use, more properties like Duval Sierrita (16b ft diam by 19 ft long) and Bougainville (18 x 21 ft) were built with single-stage ball mill grinding. Although there are larger rod mills used, especially in the iron ore industry, the largest rod mills on copper ore are at Twin Buttes (14 x 18.5 ft) and at Gibraltar (13'A x 20 ft). Table 4 lists operating data for rod mills. If a company is listed in Table I

Jan 1, 1985

By J. C. Purcupile

In this time of increased coal demand, a faster and safer method for roof support is needed. This report outlines the preparation and testing of a machine which we feel will help answer this need. The results of the tests are encouraging, and indicate that further exploration of the method should be continued.

Jan 1, 1975

By N. Gritzuk

The area of Canada with which we are here concerned consists of Yukon Territory and Northern British Columbia. Generally it lies west of the coastal mountain range and access to the sea can be gained at only selected points which are relatively few. It is served by the Alaska Highway as well as a provincial and territorial system of roads permitting transport of mining products to Whitehorse in Yukon Territory. This town it the northern terminus of the White Pass and Yukon Route, which system provides a rail service between Whitehorse and Skagway, Alaska a distance of 110 miles, and thence to Vancouver, British Columbia, by the coastal Motor Vessel C. J. Rogers, a distance of an additional 890 miles. At present the service between Vancouver, B.C. and Whitehorse, Y. T. is on a fourteen day basis. The coastal vessel, capable of carrying 3,200 tons of cargo, was specially designed and built in 1955 for handling containerized cargo and general cargo northbound, and containerized and palletized asbestos fibre, and palletized lead and zinc concentrates southbound.

Jan 1, 1961