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By J. Ludden

By Sims Some two years ago we began investigating low coke button bituminous coal for use at our Bound Brook, New Jersey plant as a replacement for anthracite. Since August, 1957 the plant has been operating continously on this coal with the exception of a new 200,000#/hr, cyclone boiler. This study was motivated not from our desire to conduct combustion research, but for the sole purpose of reducing our steam costs, The indications were that the economics were greatly in favor of the use of the bituminous coal. The attempts to reduce the costs have developed some knowledge that has enabled this program to succeed reasonably well as will be discussed later by Mr. Ludden. The production of steam size anthracite is dependent upon either the reclamation of old culm banks or the breakage in the production of the domestic sizes. The anthracite industry has not as yet been able to cut, its mining costs to the point where the production of steam sizes has become a profitable business. This has resulted in two basic problems: (1) A constantly rising price in an attempt to have those sizes carry their share of the cost and, (2) Periodic shortages of supply when the demand is good but the market price prohibits the mining of anthracite to produce those sizes only, These conditions have made the long range forecast appear as though the price would go even higher and/or the supply at times be critically short.

Jan 1, 1958

At present, discounted cash flow analysis is generally accepted as the preferred method of valuation of mine projects in surface coal mining. One of the major reasons for that is its ability to take into account the time effect of money over the life-of-mine. As a looking-forward method, it requires forecasts be made with regard to economic, technological, technical, geological, and environmental conditions that will prevail in the future. All forecasts are inherently uncertain as they are a function of the uncertainty of all variables of the discounted cash flows analysis. This indicates that the only alternative for modeling the uncertainty in the project evaluation is the development of adequate risk models to serve the decision-making process. The paper offers a model of project cash flows at risk in strip coal mining, which is an integrated corporate risk management solution. The model is based on a stochastic approach that provides the answer to the question about how large is the deviation between stochastic cash flows and the planned value (or that used in the budget) at each time step of the discounted cash flows analysis, that is due to the uncertainty in the model variables. A solution is worked out for the allowable range of the stochastic cash flows variation that determines the probability of achieving the planned value of a mine project. The Monte Carlo simulation technique is used to model the stochastic behavior of the project variables. A case study of an opencast coal mine project is provided to illustrate the applicability of the risk model. The results obtained provide a new understanding of the mine project risk as a component of the project discount rate.

Jan 1, 2006

By Garland R. Dever

Limestone resources for industrial, constructional, and agricultural uses occur on the surface and at minable depths in several areas along the Ohio River and its navigable tributaries in Kentucky. Operable thicknesses of high-calcium and high-carbonate limestones suitable for use in cement and lime production, steel manufacture, environmental-control measures, and other special uses are documented in Ordovician and Mississippian formations by analysis of 0.3-meter (foot-by-foot) samples. Active and potential mine and quarry sites afford access to water transportation and proximity to a number of industrial and metropolitan centers. Geologic factors affecting the distribution of deposits include the Cincinnati arch, regional dip into the Appalachian and Eastern Interior basins, complex faulting in and near the Western Kentucky fluorspar mining district, and late Paleozoic erosional unconformities in northeastern Kentucky.

Jan 1, 1977

By R. Heersink, C. H. Wells

Innovations in communications and real-time information systems are providing industry opportunities to improve profitability. Real-time operating data is collected from equipment across a site or even several sites into a central server for secure PC and PDA-based Internet access. This paper includes discussion of how operating staff can use wireless devices to see equipment and processing data from anywhere throughout the mines and concentrators, and how management can view results from multiple mines, all in real time. Additional opportunities enabled through the remote access by equipment vendors or mining consultants is also presented, as well as an overview of typical benefits and current installations.

Jan 1, 2002

By F. W. DeVries, O. B. Mathre

Addition of Kastone® peroxygen compound and formaldehyde to cyanide gold extraction process effluents, containing at least 5 mg soluble copper per liter, will oxidize acid-dissociable cyanide to levels less than 0.2 mg per liter, without appreciable oxidation of thiocyanate. Most of the copper precipitates and is removed by filtration of the treated solutions. The treatment can be completed in less than 2 hours when the excess of hydrogen peroxide over the cyanide is about 75-100 mg per liter. The excess hydrogen peroxide destroys residual formaldehyde and then is catalytically decomposed by the copper-containing solids that separate, so it will not interfere with subsequent operation. A pilot plant using this technology has been operated for more than four months at rates from 1-2 gpm. Clear, filtered effluent containing less than 0.2 mg acid-dissociable cyanide per liter has been produced.

Jan 1, 1981

By J. A. Finch

The photographic measurement of bubble diameter is tedious and restricted to vessels with transparent walls and relatively low bubble concentrations. A new method for predicting bubble diameter in a flotation column has been developed and confirmed by comparison with measured mean bubble diameter (Yianatos et al., 1987) Flotation columns are not mechanically agitated. This note describes an extension of Ms method to a mechanically agitated flotation machine. A recent publication by Szatkowski (1987) provided sufficient data to calculate bubble diameter by the proposed method and compare with values determined photo- graphically. Method In a two-phase system (gas and liquid), bubble diameter (db), superficial gas rate (Vg) and liquid rate (Vl), and gas holdup (Eg) are linked by two relationships (see the nomenclature at the end of the paper for definitions of symbols). [ ] Equation 1 is the definition of slip velocity (relative gas to liquid velocity) for a countercurrent gas-liquid system, and equation 2 is an adaptation (Yianatos et al., 1987) of the expression of Masliyah (1979) for hindered settling in a multispecies system. The solution for bubble diameter is by repetitive substitution of estimates of db in equation 2 until the calculated Ubs is equal to the measured Ubs in equation 1. Results and discussion Table 1 presents the data obtained by Szatkowski (1987) from a mechanically agitated laboratory flotation machine. The machine generated bubbles by directing a gas stream into the vicinity of a disk rotating at Ngh speed (up to 8000 rpm) in the flotation slurry. The air was completely dispersed in a swarm of fine bubbles. A photographic technique was used to determine the distribution of bubble size. The mean diameter and standard deviation were then calculated. Table 1 provides all the information necessary for computing bubble diameter (Vl is assumed zero). The calculated bubble diameter is presented in Table 1 as d;. Figure 1 plots measured versus calculated bubble diameter and shows reasonable agreement between values. There is an apparent tendency to undercalculate the small bubbles and overcalculate the large ones.

Jan 1, 1987

The stability of surfactant-free foam films was studied using the thin film pressure balance (TFPB) technique. In pure water, it was not possible to produce foam films due to weak film elasticity and strong hydrophobic force. In the presence of NaCl, however, the hydrophobic force was sufficiently dampened to allow metastable films to form. At 10-5 M NaCl, the equilibrium film thickness reached a maximum of approximately 130 nm at pH 6.0-7.3. The thinner films obtained on either side of this pH range was due to the increased ionic strengths associated with the HCl and NaOH additions for pH control.

Jan 1, 2009

By H. Sis, D. Dahlin, S. Chander, C. Summers

Bench-scale selective-flocculation tests were conducted on samples of Florida and Western phosphate process tailings. Physical, mineralogical, and surface characteristics were determined for both types of samples. The flocculation tests investigated dispersant and flocculant types and doses, pH, and slurry percent solids, while other parameters were held constant. Where selectivity was achieved, phosphate minerals remained dispersed, whereas silicate minerals flocculated and settled. In single-stage tests, flocculants having low to medium anionic character effectively upgraded products at moderately elevated pH. Multiple-stage tests, and tests in which flocculants of lower molecular weight were used as site-blockers, gave improved results.

Jan 1, 2002

By Daniel W. Kappes

Testing ores for heap leach operations traditionally follows a progression of bottle roll tests, bucket and column leach tests, and very expensive large field leach tests. Enough data is now available to streamline the procedures. Data from five different operations shows that recoveries from large field leach tests can be predicted from the bucket tests. Initial recovery times in large field leaches are longer than in bucket tests- a typical response is eight days versus twenty days to the same recovery percentage; but the ultimate percent recoveries (after 60 days) are normally very close. Methods of mining and stacking can radically affect recoveries from clay-rich ores. Leaching of uncrushed ores provides special testing problems, since accurate samples smaller than several tons cannot be taken.

Jan 1, 1981

By J. P. Ducarme

Numerous fatal/nonfatal accidents involving underground workers struck by powered machinery occur yearly, with continuous miners and roof bolters involved in the majority of these accidents. In an attempt to reduce these accidents, researchers at the the National Institute for Occupational Safety and Health (NIOSH) Pittsburgh Research Laboratory conducted studies of operator interactions with the motions of continuous mining machines and roof bolter boom arms. These operators generally perform their tasks in close proximity to the equipment in confined workspaces, often employing awkward postures. Since experiments with human subjects using the actual equipment were not feasible due to safety concerns, researchers opted to conduct controlled experiments in laboratory settings. Utilizing motion capture technologies and computer simulations, the studies collected data on equipment and operator movement in a range of seam heights and working postures. This article details the results of these studies to examine operating speeds based on usage and seam height.

Jan 3, 2010

By W. H. Stahl

The continuous vacuum filtration is the solid-liquid separation process that is nowadays most frequently used in the beneficiation industry all over the world. After a development time of several decades, the various filter types such as drum filter, disk filter, plate filter and belt filter can be consideres as a sound functional design ensuring a trouble-free operation. Nevertheless, the vacuum filtration does no longer meet the present and future technological requirements. The moisture content of filter cakes is too high. Their thermal secondary drying is too expensive (coal, ores) and may bring about undesirable material changes and the filter cake conditioning with water-binding substances for dumping purposes is too costly.

Jan 1, 1983

The Rasmussen Ridge phosphate mine is owned by Agrium CPO and operated by a unit of Washington Group International. The mine, located in southeastern Idaho, received the Idaho Inter-Agency Reclamation Award from the Idaho Department of Lands. The phosphate mine, located northeast of Soda Springs, was honored for “Excellence in Ongoing Operations, Soft Rock Mines over 75Acres.” Charles Ross is Agrium CPO’s plant manager. He said, “We are honored to receive this award. This recognizes the commitment and dedication of our employees toward the protection of the environment. I am really proud of the people that work at the mine.”

Jan 1, 2005

By J. C. Haptonstall

The Bismark Mine is located in northern Chihuahua, Mexico, 150 km (95 miles) southwest of Ciudad Juarez (Fig.1). The ore deposit was discovered in 1979 and was placed in production in 1992. The production rate is about 2.4 kt/d (2650 stpd) from an ore reserve of 8.5 Mt (9.3 million st). Grades currently average about 8% zinc, 0.5% lead, 0.2% copper and 50 g/t (1.46 oz/st) silver. Substantial inflows of water into the underground mine complicate the operation. Complex metallurgy poses another challenge for the operators. Bismark is a joint venture with 90% participation by Minas Penoles, SA de CR of Mexico City and 10% participation by Promociones Industriales Banamex, S A de C V (PIB S A) of Mexico City. Anew company, Minera Bismark, SA de CV was formed to operate the property. Cyprus Minerals was a 40% participant in the joint venture but, in early 1993, sold its share to Minas Penoles. The Bismark Mine is one of the largest new underground mining projects built in Mexico during the last 10 years. It is the country's fifth largest underground metal mine and provides about 1 % of the world's zinc supply at about 50 t/a (55,000 stpy). Bismark is a trackless operation with overall productivities of 7.6 t/manshift (8.4 st/manshift). The 2.4-kt/d (2650-stpd) concentrator started up in March 1992.

Jan 1, 1994

By Zephaniah Varley, Dominic Parmantier, Paul Schmall, Richard Martin, Red Robinson

Sequential excavation of the Beacon Hill Station has been accomplished in highly complex and inter-layered sand, silt, and clay soils with multiple perched groundwater tables. Successful excavation has required the installation of a deep dewatering well system used to depressurize sand layers above and within the excavated tunnel faces. In addition, the presence of sand layers in the crown of the Platform Tunnels required jet grouting at depths of up to 49m (160-ft). This paper discusses the design intent, construction, and effectiveness of the dewatering system and the selective jet grouting of sand horizons and shaft break outs for the successful construction of the deepest SEM station excavation ever completed in the United States.

Jan 1, 2007

By Tom Ireland, Tony Rock

The proposed 1.83km long twin bore A3 Hindhead road tunnel planned for contruction in 2008 is being delivered under a Highways Agency Early Contractor Involvement (ECI) contract. The proposed construction methodology is progressive mechanical excavation of a horseshoe shaped tunnel with shotcrete installed close behind the face using robotic spraying equipment. The tunneling medium is a ‘soft rock’ sandstone for which there was little previous tunneling experience, requiring extensive geotechnical investigations and an innovative approach to lining design. A 240m long portal section though sandy material utilising a steel pipe umbrella is also required. This paper describes the development of the tunnel planning and design including the most recent developments in UK road tunnels, and the innovation included in the ‘permanent’ primary lining design.

Jan 1, 2007

By W. G. Pariseau

Conventional barrier pillars in underground coal mines are intended to protect mains and sub-main entries for relatively long periods of time. Inter-panel barrier pillars that may be used in especially deep coal mines, say, below 800 m (2,400 ft), are intended to protect longwall panel faces and adjacent panel entries for the life of nearby panels. Barrier pillar failures have serious consequences, especially when failure is sudden and catastrophic. Design of barrier pillars is thus of considerable importance to safety and efficiency of operations. Locally, failure occurs at the limit to elasticity when stress tends to exceed strength. Strength of rock is stress dependent, so an analysis of stress is central to rational design of barrier pillars. There are several approaches to stress analysis for coal mine pillar design that range from the simplified but popular tributary area method including variations that use empirical definitions of the tributary area (volume) to more sophisticated numerical methods such as the boundary element method and the finite element method. Despite ease of use, empirical methods are limited and always suspect when applied to conditions beyond their scope of definition, while boundary element methods do not generally include pillars in the solution domain. The finite element method is not so restricted and thus is a natural choice for conventional and inter-panel barrier pillar safety analyses. Caving of roof strata certainly affects the distribution of stress about longwall faces and nearby barrier pillars. Conventional wisdom that considers roof strata to respond as cantilever beams indicates the longer the span the greater are loads on longwall faces and face supports. Failure of roof beams, caving, shortens spans and relieves face and support loads. The same view is applicable to barrier pillars in that long roof spans outside of barrier pillars leads to higher pillar loads than short spans. Hence, caving should be beneficial to barrier pillar safety. However, the benefit may be small depending on the extent of caving and only in the form of a slight reduction in overburden weight transmitted to the barrier pillar, especially in case of inter-panel barrier pillars. FINITE ELEMENT MODEL Finite element models for stress analysis in rock mechanics are formulated in consideration of physical laws, kinematics, and material laws. These models meet the essential requirements for equilibrium of stress, satisfy strain displacements relations, and in a general sense, relate stress to strain through material laws, often referred to as constitutive relations. The last may be complex and replete with nonlinearities. A de facto engineering material law is the elastic model in the form of Hooke?s law for linear, homogeneous materials that may be anisotropic. A limit to elasticity is imposed by strength in the form of a failure criterion or yield function that accommodates triaxial stress states. Examples used in rock mechanics are the well-known Mohr-Coulomb (MC)criterion, the popular Hoek-Brown (HB) criterion, and the Drucker-Prager (DP) criterion. The first is linear in stress, while the second is nonlinear. Both do not include any effect of the intermediate principal stress. The last is also linear in stress but includes the intermediate principal stress. All show an increase in the major principal stress at failure with confining pressure. In the UT2/3 finite element programs, the failure criterion Y is a nonlinear, anisotropic form and is also the plastic potential [1]: [YJI=+ 11NYJI=+- (1) ] where N is an exponent, usually 2, and J2 and I1 are nonlinear anisotropic forms of the second invariant of deviatoric stress and the first invariant of principal stress. The form (1) reduces to DP when N=1and isotropy prevails. When Y = 1, the limit to elasticity is reached and plastic deformation is possible. The material constants that enter (1) are all computed from unconfined compressive, tensile, and, in case of anisotropy, shear strengths. These latter strengths may be estimated from conventional laboratory test data. Of course, joints in field-scale rock masses induce different values. How different is often a matter of conjecture. For comparative studies, laboratory rock properties seem adequate. Inelastic, that is, plastic behavior occurs with strain beyond the elastic limit. Strain hardening, ideal plasticity, and strain softening are possibilities. Ideally brittle behavior with complete loss of strength is an extreme form of strain softening [2]. Caving may also be considered as strain softening with a less drastic reduction of material properties. As intact roof strata fail, fall, and form a rubble pile below, a new material is formed with greatly reduced elastic moduli and strengths, say, 1/100th of the original intact moduli and strengths. From another viewpoint, caving is simply a natural cut and fill process. Input for stress analysis by the finite element method includes: 1. Preexcavation stress state 2. Geology (stratigraphic column) 3. Structure (major faults, joint sets) 4. Elastic moduli 5. Strengths, and specific gravities of each rock type 6. Excavation sequence Output from an analysis includes distributions of: 1. Stress 2. Strain 3. Displacement 4. Local factor of safety (element fs) The last is defined as the ratio of strength to stress and in view of the failure criterion (1), is the ratio [fsJstrengthJstress=] where the numerator is the maximum possible value and the denominator is the actual value of the square root of the second invariant of deviatoric stress. This invariant is a measure of shear stress intensity but also reduces to uniaxial tensile and compressive strengths as the case may be. An fs greater than 1 indicates an elastic state of element stress; a value of 1 indicates the limit to

Jan 1, 2009

By Larry D. Wayland

Almost 100 sedimentation ponds were conceptually designed for a large surface mining study area in northeast Texas. An approximate procedure was developed to economically estimate construction quantities in order to predict surface water control costs. This procedure utilized site-specific empirical relationships developed from detailed analyses on a representative number of proposed sedimentation ponds. Use of these equations provided earthwork volumes, and spillway pipe lengths. The procedure developed for this study is presented along with the results of a verification analysis.

Jan 1, 1983

By William T. Barrett

It is said that Thales of Miletus, 600 B.C.,1 was aware of the attraction of lodestone toward iron and highly magnetic minerals. However, the birth of magnetic concentration followed shortly after the invention of the first electromagnet by Sturgeon in 1825. Sturgeon's magnet was described in the annals of electricity.2 The first attempts to concentrate minerals by the use of magnets date back to about 18473 but the first commercial drum-concentrator was apparently invented by R. Cook in 1849. Credit must also be given to the invention of P. Righter's machine in 1867,5 and to the highly successful Wenström separator devised in Sweden in 1883.8 Since the advent of these early separators there have been virtually hundreds of patents granted covering minor modifications of drum-type magnetic separators. The more important advances from a technological viewpoint have been the design of stable power supplies for d.c. magnets, the advent of ceramic permanent magnets which are now replacing electromagnets for the concentration of ferromagnetic minerals, and advances in general magnet and machine configurations.7 One of the problems facing both the designer and the user of magnetic separators is the evaluation of the body forces acting on particles in the magnetic field of the separator. Perhaps the most elegant method of evaluating the body force associated with a drum-type magnetic separator is that recently suggested by W. F. Brown.8 Brown measures the radial component of the field intensity one-half inch from the drum surface and fits a Fourier series in the angle about the drum axis (Figure 1). Once the coefficients of the series are known, it is then possible to calculate [ ]

Jan 1, 1970

By Michael V. Mergentime, Prakash M. Donde

The Nutley Quarry Trunk Sewer Project is a good example of a tunnel that went through difficult and complex ground. The 2 m (7ft.) diameter and 840 m (2,750 ft.)long tunnel passed through soft ground, mixed face, and hard sandstone. The soft ground and mixed face tunneling provided several challenges. During excavation of the tunnel, the Contractor encountered a number of unanticipated large boulders. In addition, the length of the mixed face portion of the tunnel increased by 100%. The Contractor tried various construction techniques to overcome the difficult ground conditions and complete the project. The tunneling progress was slower than expected. The Contractor claimed extra costs for differing site conditions. Through discussions and negotiations between the Contractor, Engineer, and Owner, all of the changed condition claims were settled before the construction of the project was completed. This paper addresses the difficult ground conditions, claims evaluation, and how they were resolved by cooperative efforts between the Contractor and the Owner/Engineer.

Jan 1, 2001

By J. Kircher

Nevada Goldfields' Aurora Mine consists of a 270 mtpd (300 tpd) precious metals mining and milling facility located 28 miles southwest of Hawthorne, Mineral County, Nevada and lies within five kilometers of the Nevada-California border. (Figure 1) Ore production presently is derived 85% from an open pit source with the balance from underground. The Aurora Mill is a compact conventional cyanide plant utilizing a carbon-in-leach extraction process followed by stripping of loaded carbon and electrowinning at Nevada Goldfields Kingston Mine, located near Austin, Nevada. Production commenced in December 1987 and annual gold production is currently 20,000 troy ounces of gold and 22,000 troy ounces of silver per annum, produced at an average cash cost of $178 per ounce of gold. Tailings handling at, Aurora was originally accomplished using a belt filter press producing a solid cake that was then transported to a final storage area. Although overall operating costs at the mine were' in the lower percentile of producing gold mines, the tailings handling portion of the operating costs were inordinately high.

Jan 1, 1989