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By Thomas A. Boyce

Introduction Although innovative coal extraction technology can be approached from many angles, this paper focuses on three: innovation in response to rising costs, to changing mining conditions, and to regulatory issues. Before exploring these three subjects, however, a basic question needs to be raised. Why Innovation? Coal is widely heralded, and quite correctly so, as the United States' "ace in the hole." It has been said that the U.S. is to coal what the Arabs are to oil. The implication is that we have nearly inexhaustible reserves - that coal is cheap, plentiful, and easy to extract. Indeed, because our current productive capacity is estimated to be 15-20% greater than demand, we have been given a false sense of security about the ease of future coal extraction. Therefore, innovative measures are being concentrated on developing technologies for the demand side of coal, including power generation, synthetic fuels, and industrial applications. What is often forgotten is that the unit cost of producing energy from these new technologies depends critically on the cost of the coal feedstock -in some cases the coal consumed accounts for more than one-third the total cost. Thus, while the developers of the end-use technologies struggle to improve the overall thermal efficiency of their process by one or two per¬centage points to make the energy produced competitive in the marketplace, they often are unaware of the risk of a 3 to 6% increase in coal costs wiping out the expensive, hard-earned results of their research. Energy planners are just beginning to realize that coal is not a uniform commodity available at low cost in infinite quantities. Coal is a heterogeneous material whose cost and quality vary widely with local conditions. To produce the massive amounts of projected coal-based energy at reasonable costs, we urgently need devel¬opments in extraction technology, as well as in end-use technology. Thus, coal extraction, coal conversion and utilization, and end -use energy consumption must be viewed as highly interdependent parts of an energy supply network (Fig. 1). And current emphasis on conversion and utilization must be balanced by more emphasis on upstream technologies, including coal extraction. Even more important, the impacts of innovation (or the lack thereof) in any one technology area in the energy supply network need to be discussed and understood by those working in each of the other areas. Why innovation in coal ex¬traction technology? Figuratively, we are all in the "coal boat" together - geologists, miners, transporters, processors, power producers, distributors, and consumers. Innovation is required from each of these sectors to overcome the obstacles that prevent greater use of our most abundant national energy resource. All other links in the coal energy supply network depend on large volumes of coal produced at reasonable costs. Declines in productivity, therefore, must be reversed; the more difficult mining conditions that will be encountered call for new methods and equipment, and regulatory issues must be resolved if the coal ex¬traction link is to be a strength rather than a weakness in the planned U.S. transition to an economy based more on coal and less on imported petroleum. Innovation Tailored to Regional Conditions Once the importance and urgency of innovative coal extraction technology has been grasped, the next step - implementing a research program - becomes quite

Jan 1, 1980

By S. A. Pirzada

A thermal plasma assisted laminar flow reactor has been used to evaluate the kinetics of converting alumina to sodium aluminate in flight. Free energy minimization calculations were used to determine experimental conditions. The experimental variables include: alumina particle size, alumina/sodium nitrate molar ratio, temperature and residence time. A mathematical model for the conversion was developed and validated. Experimental and mathematical results are presented.

Jan 1, 1996

By H. A. von Schonfeldt, H. Douglas Dahl

INTRODUCTION During the past decade, significant pro¬gress has been made in applying rock mechanics principles to mine design problems by a great number of researchers. Some of these new techniques have found their way into the engineering and planning offices of the operating mine. In spite of these advances, however, the application of rock mechanics in the day-to-day mine planning and design of our domestic coal mines is rela¬tively rare. There are a number of reasons responsi¬ble for this relatively slow progress, not the least of which is the question of how to measure the success or cost effectiveness of a sizable rock mechanics effort in an opera¬ting mine. Firstly, the properties of geologic materials generally vary over a wide range and it is, therefore, difficult to assess whether an improvement in the mining conditions is due to a design change or due to a change in the geologic conditions. A second reason for the slow acceptance of rock mechanics in engineering thinking is the fact that some of the concepts are quite difficult and that physical evidence of their validity to mine design is still rare. Thirdly, com¬munication between research, engineering and applications oriented people appears to function only intermittently. In our view, however, the most important impediment lies in the "state-of-the-art." Exactly how much can rock mechanics help the mine operator? The best that rock mechanics can do at present is determine whether one approach is better than another when weighed against some design objective (stress, strain, deformation). We cannot determine the abso¬lute best ground control design -- only what is better. If ground control is lost, the statement that "it's the best one we looked at" means very little to the harried operator. While the "state-of-the art" has advanced a great deal, it has certainly not been reduced to an engineering matter in that design hand¬books are available -- they are not. The success of a rock mechanics program, there¬fore, can be determined only over a long period of time of continuous effort.

Jan 1, 1977

By Robert E. Thurmond

With modern exploration techniques, it seems likely that an increasing number of the orebodies to be found in the future will be under alluvial fans or valley fill. As a result, a larger proportion of open pit mines will be confronted with the necessity for removal of substantial quantities of alluvial overburden; whereas in the past, for the most part, the open pit mine operator has dealt mainly with hard rock. If this proves to be the case, methods, equipment, and costs for alluvium stripping will become an increasingly important factor in open pit mine planning. Several operations are now faced with this situation, and various methods and equipment are used. At Pima Mining Company's operation the stripping ratio is in the range of 6/1 Well over half of the stripping volume is alluvium. The alluvium is removed almost entirely by carryalls hauling direct to a waste dump. Some removal has been accomplished by shovel and truck operation; more will be done by this means in the future. At American Smelting and Refining Company's neighboring .Mission Mine, nearly all alluvium removal is by large shovels and, trucks. U. S. Borax's Boron operation is accomplishing alluvium removal with carryalls but has been investigating other methods and equipment. The various operations in the Gas Hills Uranium District in Wyoming are stripping mainly with carryalls, although some small draglines are employed. The M. A. Hanna Company in the Mesabi Range employs large draglines and conveyor systems for removal of glacial fill. Nchanga Consolidated Copper Mines in Northern Rhodesia strips overburden with a large excavating wheel and conveyor system. There are no doubt other open pit operations with which the writer is not familiar, that are handling large quantities of alluvial overburden.

Jan 1, 1963

By V. F. Hollister

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

Jan 1, 1973

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 Dale S. Preece

INTRODUCTION The Waste Isolation Pilot Plant (WIPP) is a research facility located in Southeastern New Mexico. The WIPP is being developed by the U. S. Department of Energy (DOE) to demonstrate the safe storage of defense-related radioactive wastes in bedded salt. Since the WIPP is a research facility. a number of large-scale in situ experiments have been planned and are currently under construction (Munson, 1983). These experiments were designed to study various aspects of nuclear waste storage in bedded salt such as mechanical responses and creep closures of drifts, thermal response due to heated canisters, and thermally-induced fluid migration. One purpose of the WIPP is to develop and demonstrate a general predictive structural analysis capability for a bedded salt repository. One set of experiments, called the Geomechanical Evaluation (Room G), is heavily instrumented to study the creep around several rooms with different configuration. A Layout of the Geomechanical Evaluation room is shown in Figure 1. One portion of it consists of wedge shaped pillar where two drifts intersect at an angle of 7.5 degrees. The pillar can be seen at the bottom of Figure 1. One purpose of the wedge pillar experiment is to study progressive pillar failure in the tip region. Another is to determine how nonuniform pillar thickness affects creep closure of the drifts. An important aspect of the experiments is the correlation between experimental data and corresponding pretest finite element analyses of the site. The finite element simulations serve two purposes. First, the computer simulations aid in understanding the experimental results by providing calculated stress. strain and displacement fields that cannot be measured directly. Second, comparison of calculated and measured drift closures serves as a method for validating or improving the finite element models and the constitutive models employed. The wedge pillar geometry required a 3-D finite element creep calculation. Results from these 3-D calculations will be presented in this paper. A 2-D plane strain double drift model which approximates the wedge pillar geometry at a slice perpendicular to the drift has also been performed and will be compared to the 3-D results in this paper. FINITE ELEMENT COMPUTER PROGRAMS Two finite element computer programs. JAC (Biffle, 1984) and JAC3D (Biffle. 1986), were used. JAC is a 2-D finite element program developed for quasi-static analysis of non-linear solids. It employs the conjugate gradient iterative technique to obtain a solution Spatial integration is performed using a single gauss point in each four node quadrilateral element. An hourglass viscosity technique is used to control the zero energy modes that occur with single point integration. The single point integration combined with the explicit nature of the program and exploitation of CRAY-1 computer architecture results in very efficient execution J4: results were compared to results obtained with eight other 2-D structural creep computer codes in the second WIPP benchmark exercise (Morgan, 1981) JAC3D was derived from JAC to treat 3-D finite element models and has many of the same characteristics including single point integration and hourglass viscosity. These characteristics have made 3-D creep analyses more reasonable by significantly reducing computation time The first exercise of the creep capability

Jan 1, 1986

By Peter Keppler, Charles Jackson

WHY DO AN AUDIT? There is increased enforcement of environmental laws and regulations at State and Federal levels with attendant steep fines and even the possibility of jail sentences. We are all aware of environmental and public health and safety incidences (i.e., Love Canal, Bhopal) which have led to stringent laws and regulations (i.e., CERCLA, RCRA, SARA). In some cases, the courts have specified that a company institute an audit program as a condition of settlement of claims which related to lack of environmental control. The U.S.EPA issued an Environmental Auditing Policy Statement in the November 8, 1985 Federal Register (pp. 46504-46508) which encourages the use of environmental auditing by regulated entities. Elements of what EPA considers to be an effective environmental auditing program are outlined and briefly explained. These factors combine to indicate why there is a growing number of companies which have instituted an audit program. Briefly, the program can and should be used to monitor compliance with environmental regulations and corporate policy, to prevent violations, to improve policies and procedures, to raise the level of environmental awareness, to encourage communication among operations and all management levels about environmental problems, and to measure risks and control costs. AMAX Inc. has had an environmental audit program since the early 1970's. The AMAX Environmental Affairs Department, working with the Law Department, oversees and coordinates environmental audits and analyzes the findings which an audit reveals. Facilities are audited on a regular schedule; those having the greatest potential for environmental liabilities are audited at least annually; other facilities are audited less frequently. The question of who performs the AMAX audits was settled in favor of the audit team being independent of personnel from the facility being surveyed. Other methods of auditing could be the following: - Independent contractor (Appendix A) - Combination of independent contractor and in-house staff - Self-audit by site personnel The merits and demerits of who should conduct the audit can be analyzed to suit your particular needs. Next you need to decide on the procedure for the audit and the form for reporting and follow up. Environmental publications referenced in Appendix B can be helpful to you at this point. After deciding on the form of auditing, a schedule is prepared and followed to the closest extent possible. This schedule should be published and distributed to the sites so facilities know when they are being audited. This is not designed to be "a game of hide and seek." Various elements of the AMAX Audit Program are as follows: Team selection Pre-visit preparation Site visit Exit interview Report Using the audit Follow-up Team Selection The usual make-up of the team is the audit manager, the environmental specialist who normally provides assistance to the site or has knowledge about the operation, and a legal representative if his or her schedule allows. The legal representative will review the findings of the audit team if a personal visit is not timely or necessary. Size of the team can be expanded, but a minimum of the audit manager and the environmental specialist is best. Two sets of observations are better than one. The environmental specialist can also provide the rapport which is beneficial to creating the appropriate tone so that the site people will be as cooperative as possible. This is not to be an antagonistic exercise.

Jan 1, 1991

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 Ehsan Alavi, Brian Hagan, Colin LaVassar, Glen Frank, Derek Dugan, Rick Capka, Lisa Mori

"Tunnel face instability is one of the most common issues in EPB tunneling. The evaluation of an ideal face support pressure is a critical element in the design and construction phases of EPB tunneling. In this study, measured tunnel face pressures in three EPB tunneling projects (University Link U230 and U220 projects and the Northgate Link Extension N125 Project) for Sound Transit in Seattle, WA are compared with the calculated analytical methods. This comparison shows that the analytical methods generally overestimate the required tunnel face support. The best practice for managing the face pressure in EPB tunneling operation is discussed based on the experience gained through these projects. Furthermore, the algorithm that is developed for identifying ground loss in these projects is addressed. INTRODUCTION Earth pressure balanced tunneling relies on using muck within the chamber to support the face. Muck is pressurized by controlling the advance rate of the tunnel boring machine and the speed of the screw conveyor. Target face pressure is typically calculated by using the geotechnical assumptions in the geotechnical baseline report. The concept behind EPB tunneling is that the application of sufficient face support pressure will prevent tunnel face instability and the resultant over excavation. Tunnel face instabilities caused by a lack of sufficient face support can lead to ground loss and surface settlement (Anagnostou & Kovári 1996, Kirsch 2010). A certain support pressure must be achieved to counteract water and earth pressure with an adequate factor of safety. While the water pressure can be predicted accurately, the determination of the resultant earth pressure on the face is less precise. This becomes manifested in the relatively high partial safety factors on the earth pressure."

Jan 1, 2016

By A. Noble

"In coal and mineral processing, several numeric parameters are used to quantify the separation performance of unit operations. For density-based separations, the partition curve is used as the starting point for many stages of diagnostic analyses. The curve is often quantified by its midpoint slope sharpness, and this single value is subsequently used for technical evaluation, benchmarking, and comparison. While this analytical approach does depict misplacement by relative density class, it inherently prioritizes middling separation and does not consider process economics or the micro-price value of individual classes. Alternatively, the organic efficiency parameter does consider process economics; however, it does not reveal information on the root cause of lost revenue or misplacement. In this paper, a new analytical approach is introduced which reconciles the technical misplacement data derived from the partition curve with the economic data derived from micro-pricing. The result is a new analytical procedure which reflects a better use of performance data. INTRODUCTION Partition curves are commonly used in the characterization, evaluation, and simulation of various physical separators. In mineral processing, partition curve analysis is routinely applied to hydrocyclones, screens, and density separators. Graphically, the partition curve depicts the recovery of particles as a function of a physical property on which the separation is based (e.g. particle density, size, magnetic susceptibility, etc.). Given the probabilistic nature of many real separators, the partition curve often exhibits a characteristic “S” shape (Fig. 1), similar to a cumulative normal distribution. Other geometric features of the curve are also depicted in Fig. 1, including the high and low closeout (or bypass) values, the midpoint slope, and the separation cut point (property value with 50% recovery). Partition curve analysis was originally proposed to evaluate coal washability (Tromp, 1937). Given the limited computational resources at the time, the analytical methods relied heavily on graphical construction and interpretation of the curve. Owing to the simplicity of the calculation and interpretation, many of these graphical approaches are still widely used for performance analysis today. The most common calculations are based on the slope of the partition curve at the midpoint. Increased slopes correspond to enhanced separation performance, since a perfect separation exhibits a stepwise transition from 0 to 1 at the desired cut point (i.e. an infinitely large slope). One common formulation of this slope, the Ecart Probable (Ep), is derived by evaluating the slope between the 25% and 75% recovery points:"

Jan 1, 2015

By Alfredo Valdivia, Hamed Nejad, Vincent Monaco, William Cao

"The East Side Access (ESA) CM005 Contract - Manhattan South Structures - consists of the installation of the waterproofing and final concrete lining for railroad tunnels, 3-Level Wye track turnout structures, segments of the Main Cavern Station, and a series of Ventilation Structures excavated beneath the existing Metro North Railroad (MNR) Grand Central Terminal. Several key team decisions made collectively by the Construction Manager (CM), Designer, and the Contractor has allowed the scheduled construction duration of 29 months to be maintained. These strategies include the use of alternate design and construction methods, including the use of pneumatically applied concrete lining and precast lining. This paper discusses these strategies, and illustrates how successful teamwork and cooperation during the execution of a large-scale construction project can lead to lowering costs, improving quality, and maintaining the overall project schedule. INTRODUCTION The East Side Access Project The Long Island Railroad (LIRR) currently transports commuters from Long Island into Manhattan, terminating at the already over congested Penn Station on the west side of Manhattan. Once completed, the ESA Project will provide LIRR commuters direct access to the east side of Manhattan underneath Grand Central Terminal. The ESA Project will help alleviate the congestion at Penn Station, which currently accommodate the NJ Transit, Amtrak, and LIRR lines; reduce travel time for LIRR passengers traveling to the east side of Manhattan and facilitate connections to the NYCT Subway System and Metro North Rail Road. The ESA Project includes building new tunnels and facilities under Manhattan and Queens. (See Figure 1).The Manhattan Alignment In the Manhattan segment of the ESA Project the excavation in hard rock of tunnels, cross passages, caverns and shafts has been completed by tunnel boring machines, drill and blast, and road header. The construction of the structural concrete final lining, interior structures and fit-out of the previously excavated caverns and tunnels are currently being performed under Contract CM005 Manhattan South Structures awarded to Michels Corporation and Contract CM006 Manhattan North Structures awarded to Frontier Kemper Constructors Inc. This paper focuses on the design and construction aspects of Contract CM005 which is the first ESA Contract in Manhattan to successfully construct concrete final linings in pre-existing excavated areas."

Jan 1, 2016

By S. E. Restrepo

To innovate and make the operation sustainable it is necessary to identify the externalities and impacts completely. This paper presents the results of a proposed methodology to make circular economy initiatives viable in mining operations using these aspects: • Economic, safety, and environmental (HSE) risk analysis • Technology observation • Mineralogical characterization • Identification of potential waste uses • Use of government economic benefits The results of the project can be summarized as follows: • Implementation of new technologies in the aggregates process, • efficient use of water, • the valorization of waste to convert it into a by-product, • and access to tax benefits for investment savings.

Mar 2, 2022

A Mineração da Trindade (Samitri) is a major Brazilian iron ore producer. Founded in 1939, privately owned Samitri operates six mines with a total production capacity of 14.5 Mt/a (16 million stpy). The mines are located in southeastern Brazil in the state of Minas Gerais. They are operating in the iron ore quadrangle, Brazil's primary iron ore-producing region. In addition to iron ore production, Samitri has the capability to produce 350 kt/a (386,000 stpy) of ferruginous manganese ore from two mines, also located in the iron ore quadrangle. Historically, about 70% of Samitri's production of lump ore, sinter feed and pellet feed is exported. Destinations include Belgium, Luxembourg, France, Germany, Trinidad and Tobago, China, Turkey and the United States. Alegria Mine In September 1989, Samitri started a new processing plant to treat itabirite, a low-grade iron ore. The goal of the plant was to produce sinter feed (coarse fraction) and pellet feed (fine fraction). Primary separation of sinter feed from pellet feed was accomplished with conventional vibrating screens making a 0.4-mm (40-mesh) cut. The screen oversize was stockpiled as sinter feed. The undersize was deslimed with hydrocyclones and upgraded by conventional froth flotation. Resulting pellet feed grades ranged from 3% to 3.5% Si02. A change in pellet feed requirements by Samarco Mineração SA, a Samitri subsidiary, called for Si02 levels below 1 % and increased production levels. Moreover, market demands called for a change in sinter feed size distributions. Unfortunately, lower pellet feed Si02 levels could not be achieved with the existing process. This was because of the presence of 35% +0.15 mm (+100 mesh) in flotation feed. Flowsheet development. Laboratory and pilot-plant studies conducted at Samitri's research facility found that to achieve lower Si02 levels, the +0.15 mm (+100 mesh) in flotation feed must be kept below 5% to 10%.

Jan 1, 1997

By T. H. Combs, G. T. Lineberry

Increased productivity in the US underground coal industry has taxed pre-existing conveyor belt haulage systems. The usual remedies are costly and can disrupt production. A reduction in the belt feeder discharge rate on shuttle car sections results in more uniform loading and reduces spillage vulnerability. An investigation was made to determine the effect, if any, on shuttle car discharge time arising from lower feeder discharge rates. A five-section, underground mine is modeled, and belt spillage is introduced. The sectional feeder discharge rate is progressively lowered until spillage is precluded. At each level, the simulated spillage is measured and compared to the loss in face production. Methods for predicting the loss in face production from lower feeder discharge rates were investigated.

Jan 1, 1995

By G. Sanders

Discussion by G. Sanders The studies reported on in this paper were initiated to draw attention to the severe contamination problem in the Section 30 drilling program at Chimney Creek. The lithologic-subset sampling study reached a different conclusion from that presented in your paper, and I wish to comment on your subsequent analysis of the data and your conclusions. Request for more complete data In the section on subsampling, you mention that the subordinate lithologies were separated and sampled, yet only the dominantlithology gold value is plotted in Fig. 4. In a contamination study, the reader is interested in the assay values for the individual subsets. Please include a table of the subsample assay data in your reply. Also, please indicate which analytical methods were used to arrive at the gold values in the subsampling study. Turning barren rock into low-grade ore Figure 5 is very revealing and typical of all of the cross sections in Section 30. Note the long strings of low-grade mineralization spread out for hundreds of feet below the ore zones. There were some very high gold values found in certain contaminated fractions during the subset sampling. The conclusion, here, was that the distinctive, strongly-mineralized dolomite layer was probably loose and crumbly and continued to disintegrate during drilling. This caused salting of the unmineralized rock samples below. Missing the high-grade part of the ore body In your statistical analysis, you directly compare the reverse circulation assays to the diamond drill assays in Section 30. Two points argue against a direct comparison and suggest the differences are greater than the 3 % that you report. First, any core loss in a gold zone most likely means that the true gold values are greater. The drillers lost significant amounts of the clay-rich, Section-30 gold mineralization. Also, the initiated salt-mud system, an attempt to improve the core recovery, met with little success. Second, the practice of not sampling core geologically, but instead sampling on even 5-foot intervals, adds a deliberate dilution to the core assay values by including a portion of nonmineralized rock in the first and last samples of each high- grade intercept. The result is often a pair of low-grade assay values on either side of a high-grade gold zone. In reality, a high-grade gold zone has a very sharp assay wall that is often bounded by barren rock. This sampling method may make the diamond-drill core assays more like the reverse circulation values and may help explain the statistical similarities you found. However, it does not represent the true gold values in the high-grade parts of the deposit. You cannot deny that, by careful geological sampling of the drill core, higher and sharper assay values will be obtained. The low core recovery and the diamond-drill-core sampling method used act together to lower the diamond-core assay values. The 3% difference you found between the reverse-circulation and diamond-core assay values could be much larger when you consider what the true diamond-drill core values would be with optimum core recovery and a geologic sampling method for the core. Should statistics have been applied here? The statement "... that reverse circulation holes have overestimated the values of some ore zones and underestimated the values of others" (p. 345) is not correct. The subsampling confirmed what the cross sections hinted at in Section 30. Namely, beneath the high-grade zones, the reverse circulation holes created, by contamination, large intercepts of low-grade ore in regions of barren rock. Because the low-grade material was not there to begin with, this is not a process of overestimating low-grade mineralization. The next statement that "the average result is similar to that of the diamond drill holes" may apply to the data set numerically, but it is not true when viewed spatially on cross sections. Adjacent reverse circulation and diamond drill holes are almost impossible to correlate, high-grade zone values vary widely and many low-grade intercepts make no geologic sense. The subset sampling and cross sections presented in the first part of the paper show that the reverse circulation portion of the data set has some serious problems, as highlighted above, and should not have been dealt with statistically at all. Conclusion Each ore body is different, and each drilling method presents unique sampling problems. In this case, the diamond drill is the

Jan 1, 1994

By Staff

"In 2014, the estimated value of total mine production of nonfuel minerals in the United States increased. The quantity of production increased for most mineral commodities mined in the United States, as did prices. Notable exceptions were the declines in prices for most precious metals. Minerals remained fundamental to the U.S. economy, contributing to the real gross domestic product (GDP) at several levels, including mining, processing and manufacturing finished products. Following the reduction in construction activity associated with the 2008- 2009 recession that continued through 2011, the construction industry continued to expand in 2014, with increased production and consumption of cement, construction sand and gravel, crushed stone and gypsum, mineral commodities that are used almost exclusively in construction. Other segments of the nonfuel minerals industries increased as well. Trends in other sectors of the domestic economy were similar to those for industrial mineral production (Table 1).Discussion of mine production is frequently segmented according to the type of materials produced within the broad categories of metals and industrial minerals. Industrial minerals can be further subdivided as construction aggregates and other industrial minerals. The materials with the highest unit values, usually metals, tend to have very low production rates compared with those of many low-valued materials, such as crushed stone or construction sand and gravel. Therefore, for discussion and analysis of the performance of the minerals industry in general, the value of production is used rather than tonnage produced. Although the value of metals and industrial minerals mined in the United States are similar, if tonnages are compared, construction aggregates dominate, dwarfing the production of all other minerals combined.Mineral industry performanceThe estimated value of nonfuel mineral raw materials, including metals and industrial minerals, produced at mines in the United States in 2014 was $77.6 billion, a 3.5-percent increase from $75 billion in 2013 (Table 2). Net exports of mineral raw materials and old scrap were $15 billion (Fig. 2). Domestic raw materials, along with domestically recycled materials, were used to process mineral materials worth $697 billion. These mineral materials, including aluminum, brick, copper, fertilizers and steel, and net imports of processed materials (worth about $41 billion) were, in turn, consumed by downstream industries with a value added estimated to be $2,530 billion in 2014.In the United States in 2014, mine production of each of 14 mineral commodities was worth more than $1 billion. These were, in decreasing order of value, crushed stone, copper, gold, cement, construction sand and gravel, iron ore (shipped), industrial sand and gravel, molybdenum concentrates, phosphate rock, lime, salt, zinc, soda ash and clays (all types)."

Jan 1, 2014

By Douglas J. Lacke

1974 marked the start of a new era for Climax Molybdenum Co., a division of AMAX, with the initiation of an open pit operation above the existing successful underground block panel caving mine. A number of problems posed a challenge to the Climax open pit. Highly fractured rock, old mine workings, permafrost, and more than 7.6 m/a (300") of snowfall forced changes in normal operations. Final pit wall design was altered. Production bench heights were lessened due to old workings. These and other modifications contribute to a highly successful open pit operation.

Jan 1, 1981

By Güldane Atesoglu, İsmail Atilgan

The changes in the elemental contents and the structural properties of chalcopyrite concentrate when roasted were studied, and the main reactions contributing to the leaching process revisited. Firstly, TG/DTA measurements were performed to identify the roasting temperatures that influenced the type of species reacting with sulphuric acid leaching liquor. The temperatures of main reactions, whose output may affect the subsequent leaching process, were analysed for the concentrate at hand, and concluded to be generally lower than those reported in literature. Secondly, the obtained results were confirmed by images and XRD measurements on the roasted concentrate at temperatures in between 350 and 600 °C with increments of 50 °C. Finally, the leaching behaviour of the roasted specimens in a stirred sulphuric acid solution examined and the Cu/ Fe selectivity investigated that the maximum Cu/Fe ratio in the pregnant solution would nearly be achieved for the roasting temperatures lower than 500 °C, since the oxidation temperature of iron sulphate is found lower than that of copper sulphate.

Aug 11, 2022

By Thomas A. Thorson

Bentonite is a common name applied to a variety of clays that develope swelling properties when associated with water. Useful bentonite deposits occur in many Western and Southern states. Their origins may consist of the alteration of volcanic sediments or the decomposition of shallow lake deposits. The specific type of bentonite that will be discussed in this review will be only the Western or high swelling sodium type bentonite that is mined primarily in Wyoming. Wyoming bentonite first entered the market in the late 1880's when it was used as an oil absorbing cosmetic. In the 1920's it was discovered that superior binding properties were present which created new opportunities in pharmaceutical, animal feed, and steel foundries. The swelling properties that were exhibited by the high swelling Wyoming bentonites soon led to its use in oil exploration when the rotary drilling process developed a need for multipurpose fluids. Bentonites in general and Wyoming bentonite in particular have fascinated chemists and process development specialists since the 1920's. There exists a long list of industrial uses that has led to hundreds of patents and extensive literature reviews. The binding and swelling properties are developed as a result of large surface areas and high cation exchange properties. Water, being a di polar molecule, attaches to the ionic vacancies and forms a network of tightly held water particles which in turn produces an impermeable seal or a binding characteristic in unconsolidated powder. Newer uses for Wyoming bentonite that once again utilized the binding and swelling properties include the pelletizing of iron ore and a vast array of civil engineering applications including slurry trenching, lagoon sealing, and garbage landfill sealing.

Jan 1, 1983