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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 Rudolf Kvapil

INTRODUCTION Sublevel mining is a mass mining method based upon the utilization of gravity flow of the blasted ore and the caved overlying waste rock mass. As with any other mining method, sublevel caving has advantages and dis¬advantages which must be carefully considered and evaluated. The major advantages of sublevel caving are dis¬cussed as follows: Because all of the mining activities are executed in or from relatively small openings, sublevel caving is one of the safest mining methods. Drifts, which are the pri¬mary working places, are distributed in a uniform pat¬tern on all levels. Normally the maximum dimensions of the sublevel drifts are about 5 m wide and 3.7 m high. The transportation drifts can have the same section, or the height may be increased to about 4.5 m when trucks are loaded in the transport drifts. The stability and safety of such drifts in competent rock can be easily controlled by smooth blasting or by a combination of smooth blasting with shotcreting. In less competent rock masses, stability can be achieved by combined reinforc¬ing, for example, by a combination of smooth blasting, shotcreting, and rockbolting. The major mining activities can be broken down into three groups: drifting and reinforcing; ore fragmenta¬tion, i.e., production drilling and blasting; and ore draw¬ing, loading, and transportation, and all are relatively simple. Because of the repetitive nature of the mining system, one can standardize almost completely all min¬ing activities. This means that a high degree of work efficiency can be achieved. Because the components of mining production in sublevel caving can be standardized, a high degree of mechanization is possible. In modern sublevel caving the sections of drifts and tunnels are sufficiently large to allow the introduction of large trackless mining equip¬ment. The advantages of a trackless system can be then broadly utilized not only for direct mining but also for all services, including the transportation of mining per¬sonnel to the working place. The flexibility of mining is very good. Standardiza¬tion and specialization of mining activities and equip¬ment on separate levels (lower level or levels in de¬velopment, upper level or levels in production mining) together with the trackless system yield a high degree of flexibility. This allows a rapid start-up of mining and good flexibility in making production rate changes. The method lends itself to good work concentration, organization, and working conditions. Normally, on the lower levels, various phases of development are under¬way. Upper levels are in various stages of extraction. Therefore the work can be easily organized into a sys¬tem which excludes interference between mining activi¬ties. Safety of mining (in small dimension openings), good work organization, high mechanization using large modern mining equipment, etc., comprise very good working conditions. Naturally such a system enables a high work concentration and rationalization of separate specialized mining activities and therefore mining by sublevel caving can be effective and relatively in¬expensive. The major disadvantages of sublevel caving, on the other hand, are: There is a relatively high dilution of the ore by caved waste. Various types of ore loss can occur. When the ex¬traction limit (that point yielding the maximum accept¬able amount of dilution) is reached, the remaining highly diluted ore represents an ore loss. Some ore is lost in passive zones located on the level of extraction between the active zones of the gravity flow. Part of the ore from these passive zones can be recovered together with ore extraction on the lower sublevel, but some un¬diluted and often not fragmented ore located in passive zones above the plane of the footwall is lost. In gen¬eral, these losses are larger as the inclination of the ore body and the footwall is reduced. A relatively large amount of development is re¬quired. This includes transport drifts, usually located in the footwall waste rock on each sublevel, and sub¬level drifts, which connect the active mining areas to the transport drifts and as a result are partially in ore and partially in the waste rock of the footwall. The waste rock length increases as the inclination of the ore body and footwall decreases. It also includes orepasses, used for transport of the ore or waste from the separate sublevels downward to the main haulage level, and normally driven in waste; and inclined drifts or tunnels, which provide a connection for the trackless equipment between the main haulage level and the separate sublevels and are driven in waste. Finally there is the de¬struction of the surface through subsidence. To maximize the ore recovery, minimize the dilu¬tion, and achieve a high efficiency of mining by sub¬level caving, good data regarding the gravity flow pa¬rameters for the blasted ore and the caved waste are of utmost importance. The exact type and amount of data required depend upon the purpose and needs of the study. For the first feasibility study, it may be sufficient to utilize the data from other sublevel caving operations with similar conditions and circumstances. For any higher level of mine planning it is clear that more exact data, including analytical and experimental analyses up to full-scale in-situ testing, are necessary. Basic gravity flow principles and design guidelines for the application of the sublevel caving mining method are presented in the following sections. Although some¬what simplified, they should provide a basis for mine planning and operation. The gravity flow principles described can be effectively applied to other mining situations, with some modification. Also, steep dipping coal seams can be effectively mined by modified sub¬level caving.

Jan 1, 1982

By Edie Trimmer

Utah statute defines sovereign lands as "those lands lying below the ordinary high water mark of navigable bodies of water at the date of statehood, and owned by the state by virtue of its sovereignty." The lands within the bed of Great Salt Lake (GSL) are, by this definition, sovereign lands, acquired at statehood in 1896 in accordance with the "equal footing" doctrine, granting each state control and ownership of navigable waters and the lands underneath those waters within its borders. Under public trust doctrine, the state, as trustee for the people, bears responsibility for preserving and protecting the right of the public to use of the waters for navigation, commerce, fishing, recreation, and wildlife habitat. Also by statute, sovereign lands are defined as "state lands," to be managed by "multiple use sustained-yield principles." The Division of Forestry, Fire and State Lands is given management authority for sovereign lands and, as manager, has responsibility to prepare comprehensive plans, initiate studies of the lake and its resources, implement comprehensive plans through state and local entities, and coordinate the activities of various divisions within the Department of Natural Resources (DNR). The Division of Forestry, Fire and State Lands also has responsibility for management of mineral leasing on sovereign lands. The many resources on the lake--water, minerals, wildlife, recreation, archeological and historical values--are managed by as many state agencies which occasionally creates conflicts. The brines of GSL contain several ions that crystalize into valuable minerals during evaporation. The major ions in the lake are, in order of relative abundance, chloride, sodium, sulfate, magnesium, and potassium. Mineral products which are currently extracted from lake brines are sodium chloride, magnesium chloride brine which can be sold as flake magnesium chloride or further processed into magnesium and chlorine gas, and potassium sulfate. Mineral products which have potential for extraction include gypsum, sodium sulfate, and trace amounts of lithium, boron, and bromine. The GSL contained an estimated 4.3 billion short tons (st) (3.9 billion metric tons [mt]) of dissolved salts in 1998. Utah Geological Survey (UGS) estimates of the dissolved salt content in GSL have fluctuated from 4.0 to 5.5 billion st (3.6 to 5.0 billion mt) due to the dynamic conditions in the lake as salts are precipitated and redissolved, and due to the diversion of brines from GSL, such as the West Desert Pumping Project. The lake has four areas of varying salinity, separated by dikes or other man-made structures: north arm and Stansbury Bay brines at near saturation (25 to 27 percent total dissolved solids [TDS]); the main body of the south arm with concentrations ranging from 7 to 15 percent TDS as lake elevations fluctuate; the waters in Farmington Bay at approximately 3 to 5 percent TDS; and Bear River Bay at <1 to 7 percent TDS. The percent TDS in Bear River Bay fluctuates with lake level, and changes in Bear River inflow. The transfer of salts from the south arm to the north arm has raised questions about the viability of the mineral and brine shrimp industries. The UGS and the U.S. Geological Survey (USGS) continue to monitor salinities at designated sites on the lake to document changing lake salinity. A recurrent theme is that placement of dikes and diversions can have significant and rapid impacts on various conditions in the lake. Hydrocarbon resources on the lake are significant, but presently undeveloped. The hydrocarbons are low gravity (4 to 9 degree API) and tar-like, contain high nitrogen concentrations, and up to 12 percent sulfur. The unusual characteristics of the oil have been the subject of studies by chemists at Weber State University and University Louis Pasteur de Strasbourg. However, these resources are difficult, and at present, uneconomic to extract using current technology because of the nature of the hydrocarbons, and production in "an offshore, highly saline environment." Oolitic sand deposits make up many of the beaches and shorelines around the lake. Because of their high calcium carbonate content, oolites have been used by Magnesium Corporation of America (MagCorp) and its predecessors for acid neutralization and dike construction. Oolites are also used in very minor amounts in flower drying. The Utah Division of Oil, Gas and Mining reports up to 130,000 st (118,000 mt) mined annually by MagCorp from U.S. Bureau of Land Management (BLM) lands adjacent to GSL. Currently, there are twelve producing mineral leases which generated slightly more than $1,000,000 in royalties during calendar year 1998. IMC Kalium Ogden Corp. (IMC Kalium) produces potassium sulfate and magnesium chloride from brines concentrated through solar evaporation in Bear River Bay and Clyman Bay. By-product sodium chloride is transferred to IMC Salt, which packages and sells the salt. MagCorp produces magnesium metal from brines concentrated in Stansbury Bay. Cargill Salt produces sodium chloride from brines provided by MagCorp under a lease agreement. Morton Salt produces salt at the southeast end of Stansbury Island. Lastly, North Shore Limited produces cocentrated brines for use in dietary and mineral/vitamin supplements near Spring Bay in the north arm of the lake. Producers of magnesium, potash, and salt from GSL contribute significantly to the value of metals and industrial minerals in Utah. Together these companies contribute approximately $240 million in gross value, or 18 percent of the value of the state&apos;s nonfuel mineral production. Most of this production is exported.

Jan 1, 2001

By D. J. Lachel, Douglas E. Hansen, Bruce Kennedy

The very high cost of exploration and development for new mines requires that a carefully planned, complete, and technically competent geotechnical evaluation be applied to the ore body and the surrounding rock. The ground conditions must be predetermined sufficiently so that, during the development and exploitation stages, unexpected conditions will be of minor consequence. Ground conditions are a major factor affecting the cost of excavated items such as shafts and inclines. For example, shafts cost less to sink in tight dry ground than in highly pervious ground below the water table, with all other factors being equal. The proper mining methods also is dependent upon the ground conditions. Bad ground for room-and-pillar mining may be ideal for caving methods; whereas, ideal ground for room-and-pillar mining would be extremely difficult for a caving operation. To do a proper engineering design, the parameters of strength, deformability, and permeability must be determined, in addition to the loads that will be applied as a result of the mining activity. The required degree of accuracy for each parameter depends upon the individual case. For example, the strength of massive rock in which openings of small cross-sectional area are to be driven at shallow depths needs to be resolved only within several thousand kilopascals (pounds per square inch). Under such conditions, the rock strength is not likely to be of great economic importance unless the rock is very weak. Large openings at great depths require a much more accurate appraisal of rock strength and deformation characteristics. Ground water conditions are important at any depth. Today, a number of techniques exist for evaluating conditions before opening up the ground for development. These techniques and their applications are described in more detail later in this text. The evaluation of ore body ground conditions should start at an early stage, as soon as the deposit has been proven viable for a mine. Economic geology investigations mainly serve to define the value and limits of the ore deposit. To do this properly, the geology of the site is determined in at least as much detail as is necessary to define the "controls" for the ore body. The ground-condition evaluation must determine the engineering properties of the ore body and the surrounding material. This includes the structural geology as it relates to rock strength; i.e., the frequency, extent, and orientation of discontinuities such as faults, joints, bedding, etc. Ground water conditions also must be determined. It is important that the detailed geology of the site be determined at an early date. Reasonable geotechnical parameters must be established during this time period, so they can be used in mining model studies that include both physical and mathematical models. The ultimate goal bf-the geotechnical evaluation is to predict the ground behavior as the excavations are made and how the ground behavior will affect the safety and economic well-being of the mine. The formulation of a ground-condition evaluation program should be planned carefully to determine the particular geotechnical parameters only in as great a detail as will have a bearing on the mine design. For instance, it is not necessary to make numerous triaxial tests in strong rock when the mining will be conducted within a few meters (feet) of the surface in a regime of gravity loading and an absence of tectonic stresses. Similarly, it is not necessary to invest large amounts of time and effort in the determination of joint strengths if the orientations of the joints are such that the failure could not occur along them. Generally, the determination of ore body ground conditions can be divided into four main steps: 1) The general and structural geology of the site must be determined; this is determined primarily during the economic evaluation. 2) The material behavior characteristics of the deposit and the vicinity must be determined. 3) The ground water hydrology of the deposit and the vicinity must be determined. 4) Miscellaneous ground support considerations must be established. Gross features are defined first, followed by a series of increasingly detailed investigations. Specifically, the geology study should start by utilizing techniques such as remote sensing, topographic map interpretation, geologic mapping, etc. After the gross features, such as major faults, have been delineated, the studies are directed toward defining the more detailed elements of jointing, rock strength, ground water hydrology, and other characteristics. This information is used to formulate an assessment of the strength of the ore body and the surrounding rock. Finally, all of the gathered information is integrated to make an assessment of the rock behavior in relation to the types of openings, the excavation methods, the support requirements, etc., that should be used for the rock conditions that are present. The mine design and the geotechnical assessment should be an iterative process. General conclusions about the rock conditions will dictate a type of mine plan. The specific mine plan then should be modified by specific geotechnical considerations. For example, an initial assessment may conclude that the ore body is strong and that a strong rock unit is present over the ore body. Thus, one definite possibility is to mine the deposit by room-and-pillar methods. However, subsequent investigations may reveal faults through the deposit, requiring relocation of the mine openings so that pillars and other structures are in sound rock. In other words, the final mine plan should be compatible with the geotechnical aspects of the ore body.

Jan 1, 1982

By M. D. Flavel, J. C. Motz, G. V. Jergensen

Introduction Compared to the gyratory crusher, the cone crusher is character¬ized by its higher speed and a flat crushing chamber design which is intended to give a high capacity and reduction ratio for materials suitable to this type of processing. The aim is to retain material longer in the crushing chamber to do more work on material as it is being processed. A concept of how a piece of stone might flow through a secondary crushing chamber is shown in Fig. 45. The number of times material will be nipped during crushing will depend on material size, friability, and the geometry of the crushing chamber as well as its speed and eccentric throw. Cone crushers are usually specified in terms of closed-side setting (CSS) and a given quantity of material passing a particular square-mesh screen size. A criterion of product passing closed-side setting is often available. This value will vary according to the particular crusher design and the details of its applica¬tion. Essentially, cone crushers can be classified into distinct types, depending on their duty. Crusher sizes for all duties are described by terms arising out of common industry usage, such as 3-ft, 7-ft, etc., which refer to the crushing-head (mantle) diameter. At present the range of sizes available varies from approximately 2 to 10 ft diam. Weights vary in the range of 5 to 200 tons, and connected horsepowers range from 10 through 700. There are so many variations that specifics should be obtained from machine manufacturers. The crushing chambers that are fitted to various machines are also often referred to by terms arising out of common usage, such as standard for secondary crushers, and short head for tertiary crush¬ers, which refer specifically to the Symons design. Also used are more specific but equally vague descriptions, such as coarse, intermedi¬ate, and fine chambers. Some manufacturers such as Allis-Chalmers with their Hydrocone crushers refer to the model number of the machine by the feed opening and the diameter of the mantle. For example, a Model 10-84 Hydrocone crusher has an 84-in. diam man¬tle, and is capable of accepting a 10-in. diam sphere at the feed point to the crushing chamber. Rexnord and Telsmith define crushing cavi¬ties in their capacity tables as feed openings at minimum recommended discharge settings (closed-side settings). Each practice offers guidelines which should be carefully defined for the specific crushing problem. The utilization of a cone crusher depends on how well it is controlled and the features of the circuit in which it operates. Normally, it will operate in the range of 75 to 85% of the peak capacity. It is likely that industry pressures will call for a standarized de¬ applied crushing power. Experience has shown that the crushing process is more controllable if certain reduction ratios are adhered to for each stage. Technically, this is defined as that size of feed (in. or mm) of which 80% will pass, divided by the size of the product (in. or mm) which 80% will pass. Secondary Crushers These are the cone crushers which accept suitable feed size material either prepared by a primary crusher or occurring naturally, and reduce it to a size suitable for marketing in one stage, or make feed for subsequent crushing or grinding stages. Secondary crushers have feed openings of 4 to 25 in. in the larger (7-ft) models down to a 2&apos;/a to 4 in. in the smallest 24-in. models. Reduction ratios normally range between 3:1 and seldom more than 5:1. In the 7-ft models, the cone crusher typically makes a product all passing 2 in., but this is dependent on the machine design and properties of the material being treated. Secondary crushers most commonly operate in open circuit. They also are operated in closed circuit with a vibrating screen, depending upon product requirements and conditions where the reduction ratio does not cause excessive power draw and build-up of circulating load. Screening ahead of secondary crushers is generally recommended, especially where the feed contains more than 25% material smaller than the desired closed-side setting. Tertiary Crushers These are cone crushers that normally take secondary crusher product and reduce it to a marketable product or make it suitable for subsequent comminution steps. The reduction achieved is a func¬tion of the crusher design and the properties of the material to be treated. The reduction ratio is normally in the range between 1.5 and 2 to 1, and seldom more than 3 to 1. The tertiary crusher normally operates in closed circuit with a vibrating screen and makes a product smaller than all passing &apos;h or 5/8 in. Fine Crushers Sometimes these machines are called sand crushers and are called by various manufacturer brand names, such as Gyradisc, Hydrofine, V.F.C., and others. A cross section of a Symons Gyradisc crusher is shown in Fig. 46. Essentially, these crushers would be used in a fourth-stage crushing operation, but could be called on to reduce a screened feed size fraction from a secondary or primary crushing operation. This type of crusher normally receives feed no coarser than 1&apos;h in. that is scalped of all over-sized material and operates in closed circuit with a vibrating screen. For average materials, typical product is &apos;/e in. top size, although a -10 mesh product can be produced on those materials having suitable characteristics. The crusher is nor¬mally operated in a separate circuit from the main crushing plant because of the variations in output rates that are caused by varying physical properties in the feed. The reduction ratio is generally less than 2:1, and circulating loads are generally high. More interparticle crushing takes place in these machines than with conventional cone crushers and capacities can be more sensitive to changes in moisture content, feed size graduation, and other physical properties. Design Features For the various applications, crusher manufacturers normally have a range of crushers of different sizes and power ratings to select from.

Jan 1, 1985

By K. Takahashi

The methyl stretching vibrations for a series of normal and iso xanthates were determined using an infrared spectrophotometer, and the results were compared with the published information on the corresponding alcohols. The frequency shifts of methyl stretching vibrations may be attributed to the inductive effect of polar groups. The hydrogens in the shorter chain alkyl groups are somewhat polar and can be expected to interact with water molecules by hydrogen bonding. Contact angle, surface tension lowering, and solubility appear to be related to the induced polarity of the alkyl groups.

Jan 1, 1968

By Iain M. Mason

Linear meridional zones of bad roof conditions have been frequently encountered in producing collieries working the Permian age Lithgow seam in the western coalfield of New South Wales. The prediction of such zones ahead of mining is desirable for reasons of both safety and economics. In March, 1981, we carried out an experimental, three-component in-seam seismic survey at Invincible Colliery to discover whether such zones have any unambiguous seismic effect. The underground survey was the first of its kind to be attempted in an Australian mine. The data quality was excellent with the spectra broadband and clear. Polarisation analysis disclosed two modes: one compressional; the other shear horizontal. The SH model was clearly dispersive and well gound to the plane of the seam. Modal velocities decreased by up to 15% at the fracture-swarm site. There is a strong suggestion on the RLS maps obtained here that the distributed fault zone is acting as a composite scatterer of channel waves. The P-SV and SH wave tomograms obtained show, in a general way, the distribution of stress in the coal panel.

Jan 1, 1982

By Guy T. Jr. McBride

World food production is not matching the exploding increase in world-population. Since arable land is limited, ways are being sought to increase yields by turning_ to fertilizer. Phosphate, an essential fertilizer ingredient, is undergoing a period of growth unparalleled in history. The United States is at once the world&apos;s largest producer of food and of phosphate rock. Both food and fertilizer output in the United States are being enlarged to meet the demands posed by the threatening world food crisis. New phosphate deposits are being developed in Florida and North Carolina to assure the United States&apos; continued preeminence in this vital resource. Large fertilizer complexes are also being constructed with an eye to both domestic and foreign markets. The growth of the U. S. phosphate industry has been characterized by developing technology in the processing and distribution of phosphate raw materials. The trend from low analysis to high analysis materials accompanied by a change from a large number of small fertilizer plants oriented toward the marketplace to a fewer number of large plants located at the raw material source is illustrative. These changes in technology and distribution have also been important factors in the development of new mining areas in Florida, the Western States and North Carolina.

Jan 1, 1966

By M. Kane, O&apos

BHP Iron Ore initiated research programs in January of 1995 to develop long term plans for decommissioning of the waste rock material at their Mt. Whaleback operation. More than 2 billion tonnes of waste rock were deposited during the past 30 years. Ultimately, the operators will deposit a total of approximately 5 billion tonnes in waste rock dumps constructed near the open pit. The Mt. Whaleback operation is located in a semi-arid climate adjacent to Newman, Western Australia, approximately 1200km north-northeast of Perth, Western Australia. The primary research program includes the development of technology for the long term performance of the waste rock dumps with respect to vegetation, slope stability, surface runoff, erosion, and water infiltration. This paper focuses on the design and implementation of a cover system to prevent water infiltration to the underlying waste rock. A laboratory program and the soil-atmosphere modelling completed to design the horizontal field test plots are described. The construction of the horizontal surface field test plots are discussed together with the field instruments installed to monitor performance. The key components of the instrumentation program are the design and installation of large scale field lysimeters to measure the unsaturated flow of moisture into the waste rock from the base of the cover layer. The cover system is designed to maximize infiltration during wet periods for subsequent evapo-transpiration while minimizing surface runoff. The result is a near zero net infiltration to the underlying waste material. A key feature of the design is the use of pit run waste material as the cover material. Monitoring is in progress and will continue for at least two annual wet/dry cycles. The optimum cover system should minimize the net infiltration of water to the underlying waste rock while maintaining physical and ecological integrity for long term performance.

Jan 1, 1998

By James S. Browning

The Mineral Resources Institute of The University of Alabama conducted research on the removal of sulfur from coal using a combination of gravity concentration and flotation methods. The research utilized the gravity and centrifugal force capabilities of the Humphreys spiral to remove part of the pyritic sulfur and ash. Flotation of the Humphreys spiral coal concentrate removed additional sulfur and ash to yield products that conform to Federal air quality standards for sulfur oxide emissions. The clean coal products recovered by the combined process contained 7.7 to 8.0 percent ash, 0.12 to 0.16 percent pyritic sulfur, and 0.78 to 0.96 percent total sulfur. Pyritic sulfur rejection by the process was about 95 percent. In addition to obtaining good recoveries of high quality coal the tests demonstrated that the pyrite in the wastes could be isolated as enriched flotation products by the combined gravity - flotation apporach. Pyrite concentrates analyzing 25 to 33 percent sulfur were recovered. Potential benefits of the pyrite flotation step are recovery of a product of possible market value and the elimination of the pyrite from the waste ash which could be a potential acid water problem when impounded.

Jan 1, 1981

By Brian H. Kaye

Efficient design of grinding circuits for efficient production of ultrafine material requires rapid feedback of information concerning the fineness of material leaving the grinding device. Generating this information requires efficient sampling and adequate rapid characterization of the material leaving the grinding machinery. Recent developments-in electrooptic inspection of dry end wet material leaving a grinding device are critically reviewed. New methods for characterizing ultrafines are discussed. A rapid operating automated cybernetic (intelligent) permeability device for automatic control of grinding circuits using Blaine index and other fineness indices is described.

Jan 1, 1983

By J. A. Rusnak, K. M. Mohamed, M. A. Van Dyke, M. M. Sears, G. Rashed

"Researchers from the National Institute for Occupational Safety and Health (NIOSH) recently developed a coal-mass model to realistically simulate the deformation and loading behaviors of the ribs of coal pillars in underground coal mines. The scale-dependent properties (stiffness and strength) of coal material were considered in the coal-mass model. In this paper the orientation of face cleat with respect to driving direction was integrated in the coal-mass model. The controlling parameters (dimensionless coal-mass scale and fracture strains) of the modified coal-mass model were calibrated using a published case study of coal pillar strength. Numerical models with wide ranges of controlling parameters were conducted to simulate the test pillar. The best match between the numerical modeling results and the measured stresses and deformations was achieved by using a coal-mass scale of 30 and fracture plastic shear and tensile strains of 3.25% and 0.325 %, respectively. INTRODUCTION The failure of coal pillar ribs by itself are a major hazard in underground coal mines. Furthermore, failing ribs can indirectly contribute to roof and floor instability by increasing opening widths across intersections and entryways. Over the past decade, rib failures have resulted in 12 fatalities, representing 28% of the ground-fall fatalities in U.S. underground coal mines. Mining depth has always played a very important role in assessing rib stability. Gauna and Mark (2010) completed an analysis of rib fatalities that occurred between 1996 and 2010. One of their most prominent and influential findings was the identification of a relationship among fatalities, depth, and mining height. Therefore, Coal rib stability will continue to be a challenge for future deeper mines in U.S. coal basins. In the past several decades, research efforts have focused on roof support. The current rib control practice in U.S. coal mines is to select rib bolts through a trial-and-error approach (Mohamed et al., 2015a). Rib failure incidents can be can be substantially reduced or even eliminated by adopting an engineering-based coal rib design approach. The factors affecting rib stability in coal mines are numerous and mutually dependent. These factors include but not limited to mining height and overburden depth, interburden thickness for multiple-seam mining, coal strength, cleat density, entry direction with respect to cleat orientation, existence or absence of partings in the coal, percentage of extracted roof and/or floor rock, density of rib bolt support (Mohamed et al., 2015a). Moreover, it is difficult to isolate the particular effect of a specific factor on the overall rib stability. Therefore, a reliable empirical rib design approach is not achievable because it requires a significant number of case histories to be collected in order to have a reasonable range of all of the critical parameters. However, numerical modeling can be a practical tool to assist with rib support design if the modeling is realistic, which can only be achieved through calibration and validation of the numerical model against field data."

Jan 1, 2018

By Thomas E. Garnar

Mineral sand products have been produced from beach sand heavy minerals for nearly a century. These minerals have specific gravities of 3.0 or higher and are extremely resistant to weathering. The most common economic heavy minerals are ilmenite, leucoxene, rutile, zircon, stauarolite, monazite, garnet, and the aluminum-silicate minerals. Of these, ilmenite is the most abundant. In some cases a deposit may contain only one or two economic heavy minerals whereas in others five or six may be present. Heavy mineral mining began in the late eighteen hundreds when monazite was mined for its thorium content. This was followed by mining activity on the North Florida coast just south of Jacksonville Beach - where ilmenite and zircon were first separated during the WW I (1,8). Modern heavy mineral mining began in Eastern Australia during the late nineteen-thirties and spread to the United States during World War II. It is the purpose of this paper to review the commercial heavy mineral products, their sources, and their uses.

Jan 1, 1985

By Donivan L. Gordon

One of the several problems facing society today is that of disposing of its solid waste. Each year this nation spends approximately 4.5 billion dollars to collect, transport and dispose of some 360 million tons of domestic and industrial solid waste. Although considerable sums of time and money have been spent in researching various techniques for solid waste disposal, it is presently agreed that sanitary landfilling represents the most economic and universally adaptable solution to the problem of solid waste disposal. However, regardless of the factors favoring sanitary landfilling as the waste disposal method of choice, it should be recognized as carrying with it a potential for the degradation of ground- and surfacewater resources. Through an interdisciplinary effort regulations were promulgated during 1971 in the State of Iowa to govern sanitary disposal projects. Consistent with a set of rationales developed to preserve environmental quality, these regulations set forth hydrologic and geologic standards designed to protect Iowa&apos;s water resources. Minimum requirements have been outlined in the regulations through which the natural favorability of potential landfill sites may be evaluated. Natural favorability is primarily defined in terms of the hydrologic and geologic relationships between a proposed landfill site and those in the locality of the site. The regulations specify that leakage from a landfill to a subcropping aquifer must be less than 0.04 cu ft/day/sq ft at the base of the landfill. This factor is offered to quantify the maximum allowable leakage for all sites in unconsolidated sequences underlain by subcropping bedrock aquifers, Based on groundwater research in Iowa, the assumption is made that 50 feet of till will be sufficient to inhibit the downward flow of liquids from a landfill into bedrock aquifers. This assumes a difference of no greater than 40 feet in hydrostatic head between the water table in the unconsolidated sequence and that in the bedrock aquifer. Owing to the fact that bedrock aquifers in many areas of the state are mantled by less than 50 feet of till the regulations provide for site modification, through engineering, to meet the specified leakage standard, In such instances the leakage factor is evaluated by means of the generalized Darcy&apos;s Law Q = PIA where:

Jan 1, 1974

By Ihsan Berk Tulu, Ted Klemetti, Deniz Tuncay

"The Analysis of Retreat Mining Pillar Stability (ARMPS) program was developed by the National Institute for Occupational Safety and Health (NIOSH) to help the U.S. coal mining industry to design safe retreat room and pillar panels. ARMPS calculates the magnitude of the in-situ and mining-induced loads by using geometrical computations and empirical rules. In particular, the program uses the “abutment angle” concept in calculating the magnitude of the abutment load on pillars adjacent to a gob.In this paper, stress measurements from U.S. and Australian mines with different overburden geologies with varying hard rock percentages were back analyzed. The results of the analyses indicated that for depths less than 650 ft, the ARMPS empirical derivation of a 21° abutment angle was supported by the case histories; however, at depths greater than 650 ft, the abutment angle was found to be significantly less than 21°. In this paper, a new equation employing the panel width to overburden depth ratio is constructed for the calculation of accurate abutment angles for deeper mining cases.The new abutment angle equation was tested using both ARMPS2010 and LaModel for the entire case history database of ARMPS2010. The new abutment angle equation to estimate the magnitude of the mining-induced loads used together with the LaModel program was found to give good classification accuracies compared to ARMPS2010 for deep cover cases.INTRODUCTIONIn the early 1990s, the Analysis of Longwall Pillar Stability (ALPS) was introduced by Mark (1990) as a chain pillar design software and was generally accepted and used by the U.S. coal mining industry. Following the success of ALPS, the National Institute for Occupational Safety and Health (NIOSH) developed the Analysis of Retreat Mining Pillar Stability (ARMPS) program for designing retreat mining pillars using a similar approach as ALPS (Mark and Chase, 1997). The Australian mining industry also recognized the success of ALPS, and Colwell, Frith, and Mark (1999) calibrated the program to Australian conditions. The ALPS and ARMPS programs draw their strengths from the large databases that are used to calibrate them (Mark, 2009). However, following the Crandall Canyon Mine collapse in 2007, NIOSH had to reconsider the pillar design criteria used in deep-cover retreat mining (Mark, 2010). The ARMPS overburden load prediction algorithm was improved to more accurately predict the loading of narrow panels with high overburden depths by implementing the pressure arch concept, and this new version is called ARMPS2010."

Jan 1, 2019

By W. C. Morgan

The Federal Surface Mining Law enacted in 1978 is now being translated into regulations which must be followed in future coal mining. Substantial change to mining procedures is required to maintain stricter air and water control while mining and to return the mined out land to equal or better use than before. These changes do not come easy; and in most cases, they create a substantial additional cost. Based on field observations and mining studies, Caterpillar has been making suggestions on the use of their types of equipment to conform to the new standards at the least cost. This paper will be a brief review of some of the overburden stripping and reclamation methods suggested by Caterpillar. A production and cost comparison of dragline, scraper/dozer and truck, loader systems in 20 to 150 ft. Overburden conditions illustrates potential cost savings using mobile equipment.

Jan 1, 1980

By Robert L. Somerville

As more phosphate deposits are discovered and new mines are developed, the necessity of critically evaluating various rocks for phosphoric acid production recurs ever more frequently. This need is accentuated by the increasing number of, fertilizer plants being constructed on deep water at various locations to serve world markets and to take advantage of locally-available sulfur or sulfur-bearing ores. A major proportion of all phosphate rock mined is consumed by wet-process plants and this acid in turn is consumed, usually captively, in the manufacture of fertilizer. It is apparent, therefore, that the first question asked by the developer of a new deposit is - How will it work in a wet-process plant? and conversely, when developing a deep water, project, one of the early studies required is - Which of the available rocks shall we use? What is the correct price to pay? How do we compare one rock with another? The economic importance of even minor aspects of these questions

Jan 1, 1966

By Jonathan K. Hrica, Brianna M. Eiter

The ability to identify hazards during a workplace examination is a critical skill for mineworkers to have in order to maintain a safe workplace. While research suggests that being able to successfully recognize hazards requires a complex set of competencies, it is unclear which competencies are most critical to adequately perform a workplace examination. This paper presents a hazard recognition competencies framework as an explanation of the competencies that affect mineworker hazard recognition. To better understand how the industry is selecting and preparing workplace examiners, researchers fromthe National Institute for Occupational Safety and Health (NIOSH) interviewed nine health and safety (H&S) professionals at surface stone, sand, and gravel mine sites throughout the USA. A primary purpose of these interviews was to verify the competencies included in the hazard recognition competencies framework and to understand why these competencies are critical. Finally, the results of this study are presented within the context of current NIOSH research findings related to hazard recognition.

Jul 27, 2020

By MAURICE SUNKPAL, Taghi Sherizadeh, Kutay E. Karadeniz, Dogukan Guner, Samuel Nowak

Despite numerous empirical, numerical, and hybrid studies conducted on the subject of coal rib failure, mine fatalities related to rib failures continue to occur at around 1.5 per year between 2009 and 2021. To date, a standard methodology for the design of coal rib supports that can accommodate the wide range of unique conditions found in coal mines in the USA has not been developed, forcing mine operators to rely on the trial-and-error process or industrial legacy practices for the design of coal rib support patterns. In this study, to better understand the support systems’ effectiveness, various rib bolting scenarios were investigated. A coal mass constitutive model is implemented to a distinct element method (DEM) code, and rib factor-of-safety (RibFOS) analyses were performed by applying strength-reduction for the coal matrix and explicitly modeled face cleats. Ribs with various overburden depths and mining heights composed of solid Banded-Bright coal (BBC) were considered in this study. Different scenarios such as rib bolt length and the number of bolts were studied to quantify the effect of primary rib support density on rib stability. As a result of the unsupported coal rib analyses, a decaying exponential relationship is proposed between the mining height-overburden depth and the RibFOS. The developed coal rib models clearly show the coal-rib bolt interaction and illustrate how the addition of rib bolts restricts the movement of the coal rib. The results reveal that the strength reduction method in DEM code with the coal-mass model can aid in the development of practical primary support charts, which can help in the rib support design process. In addition, an unsupported RibFOS cutoff is proposed, beyond which the addition of primary support cannot stabilize the failed rib structure.

Aug 5, 2023

By Fengyu Ren, Kunpeng Yu, Gideon Chitomb, Qiangqiang Gao, Liqiang Ma, Jiangtao Zhai, Ichhuy Ngo

The penetration of rock fragments of overlying strata causes advanced ore dilution during drawing. The results of numerical and physical experiments showed that the proposed numerical simulation was applicable and the microscopic parameters of the model were reliable. The penetration under the variables of particle diameter ratio, density ratio, overlying strata thickness, and particle morphology was clarified. The penetration was quantitatively characterized by the discharge amount and the dilution rate. The experimental results showed that the diameter ratio and density ratio had a significant influence on the penetration. On the irregular shape of particles, the penetration is weakened. The overlying strata thickness had little effect on the penetration. This study could provide a theoretical basis for the full section uniform drawing, avoiding advanced ore dilution and further optimization of recovery rate.

Nov 13, 2023