Search Documents

By J. T. Graves

Bauxite, dolomite, and magnesite are natural occurring minerals that are used extensively in the manufacture of refractory products. In addition, magnesia produced from seawater or brine finds its way into the production of refractories. All of these materials must be processed at high temperatures, 1500°C to 2200°C, to produce an intermediate raw material (refractory grain) , which can then be used in the final manufacture of refractory bricks and shapes. During the last 20 years, the major supply of refractory bauxite has been from Guyana, and at times the annual production has ranged up to 1,000,000 tons per year. More recently, China has become a major source of refractory grade bauxite by supplying over 300,000 TPY. Other sources of bauxite in lesser quantities have been Surinam, Brazil, and the U.S.A. During 1980, the estimated available world tonnage amounted to slightly more than 1,300,000 tons. The bulk of this material has been calcined in rotary kilns except for some of the Chinese production, which was treated in shaft kilns. Recent work has indicated that an excellent product can be produced in the Nichols Shaft Kiln using much less fuel than the conventional rotary kiln. Details of the Nichols Shaft Kiln will be provided later in this paper.

Jan 1, 1982

By James C. Bradbury

Tripoli is a naturally occurring, finely divided form of silica found chiefly in midwestern and southeastern states and used commercially in fillers and abrasives. The name comes from the location of a silica deposit in North Africa and was applied in the latter part of the 19th century to similar material near Seneca, Missouri. Subsequently, the North African material was found to be diatomite, composed of siliceous skeletons of microscopic marine plants called diatoms, and the material's name was changed to tripolite. The name, tripoli, remained with the Missouri silica, which is inorganic in origin, and tripoli is now used to designate all naturally occurring silica that is similar in characteristics and origin to the Missouri material. Just to confuse things further, in commercial trade tripoli usually designates silica from the Missouri-Oklahoma field, whereas silica from southern Illinois, the other major producing area, is called amorphous silica. Each term, however, may be used commercially for any material, regardless of place of origin that has the characteristics or uses ascribed to material from either major producing area. For example, Engineering and Mining Journal routinely lists under "tripoli," prices "f.o.b. Elco, I11." and "f.o.b. Seneca, Mo. and Rogers, Ark.", and under "silica, amorphous," prices "f.o.b. I11 and "f.o.b. Dierks, Ark." "Soft silica" is another term that has been applied to tripoli from Illinois, west central Arkansas, and Tennessee, presumably to distinguish it from the more compact Missouri material. It has also been used in the ceramic trade to distinguish tripoli from "hard" silica, or silica flour, which is produced by pulverizing silica sand. Tripoli is produced mainly in extreme southwestern Illinois, in the Missouri-Oklahoma district in Newton County, Missouri, and Ottawa County, Oklahoma; and in the novaculite belt of west central Arkansas. Some tripoli is produced for local and specialized uses near Austin, Texas.

Jan 1, 1980

By G. W. Mannard

The large massive sulphide deposits present a formidable challenge to the economic geologist. They represent such huge concentrations of iron, sulphur, base metals and precious metals that it is difficult to conceive how they were derived from the crust or mantle of the Earth. The problem of the origin of massive sulphide deposits has generated a geological controversy which has lasted two hundred years and shows no sign of abating. The economic incentive to solve this problem is great, as the massive sulphide bodies are the moat important source of the world's zinc, lead, and silver, as well as yielding large amounts of copper and gold. Using the criteria of abundance, size, age distribution and gross value of contained metals, the predominant sub-group of the massive sulphide deposits is that which comprises the crudely conformable pyritic (or pyrrhotitic) deposits which occur in stratified volcanic and sedimentary rocks or their metamorphosed equivalents. Examples of this sub-group are: the Buchans, Bathurst, Noranda, Kidd Creek, Flin Plon and Sullivan deposits of Canada; the Mt. Isa and Broken Hill orebodies of Australia; the Cyprus, Rio Tinto, Meggen and Rammelsberg deposits of Europe; Ducktown and the United Verde in United States; and the Kuroko deposits of Japan.

Jan 1, 1973

By Ramkrishna Agarwal

A low grade iron ore is mined underground at Grace Mine using the panel caving method. Most of the development work is located in footwall. The runways for LHD equipments are heavily supported with steel sets on five foot centers. The sets are embedded in two to three feet of concrete. The initial cost of support is one of the largest single costs of production. During the process of ore-win¬ning, supports frequently fail, necessitating expensive repair. There is no systematic pattern to the failure of supports. A rock mechanics program was launched to isolate causes and their relative contributions to the failure of supports. The preliminary results are presented. A typical portion of a runway is instrumented to monitor rock mass movements around the runway and the corresponding response of supports as various mining activities are carried out in the neighborhood. We believe that such a program is a prerequisite to any remedial measure to alleviate pressure problem at Grace Mine.

Jan 1, 1973

By A. B. Fly

This paper presents the equipment and techniques developed to mine subsurface mineral deposits through 16 inch bore holes entirely by hydraulic methods. Use of this equipment for excavating large cavities in oil and water well producing zones is discussed. Results of surface experiments using high pressure jet streams of water to determine the cutting rates of materials are outlined. Various basic shapes of subsurface excavations that can be mined with the equipment are illustrated and discussed. An example of the cost of mining by this method is presented. Problems encountered in the field operations led to the development of a new explosion type slurry pumping system and related slurry handling equipment.

Jan 1, 1969

By James W. Collins

The bucketwheel stacker/reclaimer is suited for a variety of bulk material handling systems, especially those operating with stockpiles in excess of 30,000 tons. The bucketwheel stacker/reclaimer, although relatively new in the United States (first introduced here in 1964), has been successfully used in European countries since the early 1940's. Basically the machine is the outgrowth of the bucketwheel earth excavator and the traveling tripper boom type stacker. Since their introduction here they have achieved wide acceptance, especially with coal fired power plants. (Figures 1, 2, and 3.) The stacker reclaimer has been utilized in barge and ship loading/unloading facilities, and in the material handling systems for steel mills, where it has been used to stack and reclaim iron ore at rates up to 6000 tph stacking and 4000 tph reclaiming. They have also been employed at mine mouth power plants for the handling and storing of the plant fuel supply and to help balance the mining rate with the fuel consumption rate.

Jan 1, 1973

By Charles W. Merrill

The price records of 7 representative mineral commodities for the 68-year period 1890 through 1957 have been compiled and analyzed for significant trends. When reduced to real prices in terms of dollars of constant purchasing power or to the purchasing power of industrial wages at average rates, a substantial overall fall in prices is revealed. This downtrend contradicts the widely held concept that, heavy drafts on a mineral resource must lead to scarcity, reflected in rising prices. Three metals (aluminum, copper, and pig iron), two fuels (bituminous coal and petroleum)', and two nonmetals (sulfur and cement) have been chosen because of their preeminence in their respective categories, their significance in an industrial economy, and the ready availability of their price records. It might be added that these seven commodities. were selected before any price figures were compiled, none was selected or rejected to substantiate any preconceived notions as to price trends, The overall importance of the seven is demonstrated by the fact that, taken together, they composed over three-fourths of the value of all minerals produced in the United States in 1957. The first step in the analysis was to reduce the price records to a basis for significant comparisons, Two such comparisons have been made: (1) The quantities of each of the commodities that could have been purchased for an average hour's wage in each year; and (2) the unit price of each commodity through hhe years in terms of deflated dollars. These data are set forth in the accompanying table and two charts.

Jan 1, 1959

By L. Barry Phelps

Selective placement of toxic mine spoil has been advocated as a means of reducing environmental pollution at the source. However, methods of burial to control environmental damage are not well understood. Using the finite element method, this paper investigates the burial of toxic material within low permeability packages below the water table. The effects of differing package and spoil permeabilities, anisotropy, and infiltration are modeled to provide information about their effect on the flow system. Projection can then be made which will assist in the design of the toxic spoil packages to limit the amount of groundwater flowing through them.

Jan 1, 1983

By R. J. Smith

Tar sand deposits in Utah represent more than 25 billion barrels of in-place bitumen, 96% of the known U.S. reserves. The technological development of a hot water processing strategy for the Canadian tar sands and the commercial Production of synthetic crude oil from this natural resource suggest that similar success may be possible for the utilization of Utah tar sands to provide our country with a new domestic source of synthetic crude oil. In this regard an active research program to develop a hot water processing strategy for Utah tar sands has been sustained at the University of Utah for the past five years. The hot water process for Utah tar sands differs significantly from that used for Canadian, tar sands due to the inherent differences in the respective bitumen viscosities and the nature of the bitumen-sand association. These differences have led to the identification of certain design criteria necessary to achieve satisfactory phase disengagement during digestion. (8,9) Subsequent flotation of the digested tar sand is a subject of current attention and is considered in this contribution for the Asphalt Ridge tar sand deposit. Although contact angle measurements of pure Asphalt Ridge bitumen indicate moderate hydrophobicity, air bubble attachment to the bitumen concentrate taken from a 37 liter flotation cell is not possible. This surprising result suggests that the flotation separation is dependent on air bubble entrapment rather than on attachment due to surface hydrophobicity. The occlusion of air bubbles in the bitumen is apparent from visual examination of the concentrate, especially at lower temperatures. It seems that a bitumen bubble agglomerate forms in the impeller region of the flotation cell, the effective density of which is such to allow the agglomerate to float to the surface. A factorial design of the major operating variables in the flotation separation indicated that the quality of separation was significantly dependent on the flotation temperature and to a lesser extent on the degree of agitation. For a flotation temperature of 77C recovery of 96.7 percent was realized at a grade of 61.0 percent bitumen. The improved separation at higher flotation temperatures was found to be due to the decrease in bitumen viscosity resulting in more effective rejection of coarse sand from the concentrate.

Jan 1, 1980

By Reed S. C. Rogers

The breakage parameters of phosphate ore were determined by batch first-order grinding studies in a laboratory ball mill according to the procedures described in the text by Austin, Klimpel and Luckie, using a slurry density of 70 weight percent. The results were scaled to five larger mills of 0.56 m diameter by 0.91 m length (2 ft. by 3 ft.), 0.82 m by 1.52 m (3 ft. by 5 ft.), 3.36 m by 7.90 m (11.5 ft. by 25 ft.), 4.41 m by 9.10 m (14.5 ft. by 30 ft.), and 5 m by 11.1 m (17 ft. by 36 ft.), assuming equilibrium ball charges for the full-scale mills. The results were used to predict product size distribustions and capacities for batch grinding in the 0.56 m and 0.82 m diameter mills and open circuit grinding in the 0.82 m diameter and larger mills using a one-large/two-small reactors in series model fitted to the residence time distribution measured in the 4.41 m diameter mill. The simulated results were compared to plant data for ranges of mill capacities. It was found that the simulated product size distributions were in good agreement with the experimental values below the 70% passing sizes, but did not agree so well in the top sizes. The predicted capacities were in approximate agreement with experimental values for the pilot-scale mills, were about 7 percent too high for the 3.36 m and 4.41 m diameter mills and about 25 percent high for the 5 m diameter mill. Use of the Bond method with the measured Work Index of the phosphate ore gave capacity predictions that were 45 to 150 percent too high.

Jan 1, 1985

By David K. Sitler

Continuing advances in manufactured diamond materials used in rock cutting tools have created a new generation of high production tools for mining. This paper will review the three major manufactured diamond products available for the mining industry as supplied by General Electric Company, Specialty Materials Department, and examine the ways in which these materials are employed. As an integral part of this examination, the economic impact of manufactured diamond tools will be evaluated. Time and again, one will find that the premium price of these tools is economically justified when compared to traditional tools.

Jan 1, 1985

By Jing G. Yen

When a solid-liquid mixture is conveyed through a pipe,' four hydrodynamic conditions may develop depending upon the properties of the transporting medium, the properties of the transported material, the properties of the pipeline and the flow characteristics. They are classified as homogeneous flow, heterogeneous flow, intermediate flow and saltation. Heterogeneous flow, which is the subject of the present analysis, occurs when solid particles are coarse, of high density, and the mean velocity of the flow is relatively low, such that the combination allows the establishment of a pronounced concentration gradient in the cross-section of the pipe. In this type of system the suspended matter does not change the rheological properties of the conveying fluid, and solid and fluid phases behave somewhat separately. In recent years much work has been directed toward gaining more insight into the hydrodynamics of heterogeneous flow (1,2,3,4,5,) However, almost all knowledge obtained is empirical in nature and limited in scope. Much effort has been extended to collect experimental information on the energy losses associated with systems of slurry flow (6-13) The work of Durand and Condolios(14,15) represents one of the most extensive experimentations in this field. Through systematic test runs, the effects of the major flow-parameters were studied one at a time. From the results, they concluded empirically that:

Jan 1, 1969

By L. A. Garfield

Since the start of mining operations at White Pine in 1953, over 22 miles of drifts have been driven for the purpose of developing the ore body. The drifts, normally driven in multiples of three or five headings, serve as conveyor belt routes, ventilation ducts, equipment and haulage roads, and initial access to mining blocks. Development openings in the flat-lying, bedded ore body are driven in rectangular section from 16 to 32 feet wide and from 8. 5 to 16 feet high. The floor level normally coincides with the bottom limit of the chalcocite values and the roof or back is carried at the most competent stratum adjacent to the top of the ore. A development front consisting of five headings and associated crosscuts is advanced at an average rate of about 100 feet per month using conventional drill and blast techniques and equipment. The cyclic nature of the functions involved in driving the openings and the need to include them in production schedules preclude significant improvements in advance rates using current drill, blast, muck, and haul sequencing.

Jan 1, 1969

By Albin E. Anderson

The PSM System has been available to the mining industry for nine years. During this time, many of the more than 160 Systems sold have been put to practical use and have provided the owners a handsome return on investment. In this paper, the author will briefly discuss how the System works and how it can be put to use in automatic grinding circuit control. Throughout, detailed consideration will be given to the many factors that can significantly influence how successful a specific application is.

Jan 1, 1979

By T. J. Close

On September 10, 1991, Secretary of the Department of Interior Manuel Lujan deleted the Silver Peak Range North Wilderness Study Area from the list of U.S. Bureau of Land Management lands in Nevada to be considered for wilderness designation. This decision was based in part on the U.S. Bureau of Mines mineral investigation of the wilderness study area which found that it has indications of precious metals resources. The following is a summary of the results of that mineral investigation. The study area lies between the Silver Peak and Dyer mining districts of west-central Nevada. The Silver Peak mining district, to the east, has a long mining history dating back to 1907 when precious-metals-bearing veins were discovered. The major producer in the Silver Peak district was Sunshine Mining Company's Sixteen-to-One silver mine. During the Bureau's study the Sixteen-to-One mine was operating; it has since been idled due to low silver prices. Precious metals-bearing veins were discovered in the Dyer district, to the west of the study area, in the 1860's. Mineral activity in Dyer district consisted mainly of prospecting. The Silver Peak study area itself has no mining history and contains only a few mining claims, but the Bureau examined 13 mineralized sites, most previously unknown and unclaimed. These sites are along two zones that cross the study area from the Silver Peak to the Dyer mining district. The two zones transect a volcanic center of fractured and altered rocks that contain epithermal veins of quartz, calcite, and jasperoid with gold, silver, and base metals. At the surface these veins are poorly defined and poorly mineralized. The more than forty veins examined in the study area are as thick as 17 feet and as long as 1,000 feet. Samples from surface exposures of these veins contained as much as 0.02 ounce per ton gold, 2.1 ounces per ton silver, 1.0 percent copper, 2.6 percent lead, and 0.4 percent zinc. Analyses of the Bureau data disclosed no vein large enough or of high enough grade to be classified a mineral resource. However, most do have indications of silver-and gold-bearing resources at depth. Subsurface exploration had to be done on the vein at the Sixteen-to-One mine to outline reserves (a vein averaging about 22 feet thick and containing 6-to 7-ounces per ton silver); the same is true of veins in the Silver Peak Range North Wilderness Study Area.

Jan 1, 1993

By F. M. Stephens

This paper is to be presented at the annual meeting of the American institute of Mining and Metallurgical Engineers, San Francisco, February 1959. Permission is hereby given to publish with appropriate acknowledgements, excerpts or summaries not to exceed one-fourth of the entire text of the paper. Permission to print in more extended form subsequent to publication by the Institute must be obtained from the secretary of the Institute. If and when this paper is published by the Institute, it may embody certain changes made by agreement between the Technical Publications Committee and the authors, so that the form in which it appears here is not necessarily that in which it may be published later.

Jan 1, 1959

Well designed and maintained haul roads are the key to minimizing truck haulage on-road hazards and costs, as well as increasing productivity. However, practically designing and managing a haul road for optimal performance is often difficult to achieve. Poorly designed and built roads exhibit high rolling resistance - a 1% increase in road rolling resistance can typically reduce speed on ramp by as much as 10% - and on the flat by up to 26%. In the current economic climate, investment and operating decisions come under scrutiny. In the long run, this scrutiny returns improved efficiencies and leaner, healthier operations. The focus of this evolving evaluation process should and will certainly fall on haulage operations ? simply by virtue of their contribution to overall cost of operations ? often in excess of 50% of total costs for deep open-pit mines. Whilst the end result - improved efficiency and reduced cost per ton hauled ? is not in itself problematic, it is the route, or process followed to achieve these savings that needs to be carefully managed. We can be guided on this journey by our understanding of how a road design is developed, and, critically, the interplay between a good design and safe, cost efficient haulage. This paper briefly summarizes the evolution of mine haul road design, from the seminal USBM work of Kaufman and Ault in 1977, through to current geometric, structural, functional and maintenance management design components. These augmented design and management guidelines have been developed over the past decade, both in response to the requirements of mine operators for more safe and efficient haulage systems, and the truck manufacturers? requirements for a more predictable and controlled operating environment. These developments have been paralleled by the need to minimize haulage hazards, both from a health and safety perspective. Whilst improved mine haul road design does indeed reduce haulage accidents, recognition also needs to be given to human factors which are a significant contributor to haulage accidents. The human factor is the most problematic to address in a road design. It is often easier break the link between the interactive effects which may lead to accidents than trying to predict and reduce human error. These human factor interactive effects include the geometric, structural and functional design components and to prevent an accident or reduce the severity of its consequences, a road should be more accommodating to human error. The more that is known about human error, the better a road can be designed to accommodate those actions or non-standard practices that would, on a poorly designed road, invariably escalate an error into an accident.

Jan 1, 2010

By Jorgen Elbrond

The calculation of open pit operations' capacities has been the subject of many studies using computer simulations. The common problems of simulation are that if one desires a realistic picture of an operation the programming work becomes long and expensive and that the program requires a large computer and long time to run. The method presented in this paper overcomes these problems by using a straightforward calculation procedure. It is based on the queuing theory's simple formula for waiting time in a closed circuit and uses a system of correction factors, which reflect the loading, travel, and dumping times' relative variabilities and which have been found by 1) mathematical analysis and 2) simulations of basic circuits. The method includes the main parameters of the operation: numbers, sizes and operational availabilities of trucks, shovels and dumping point, distribution of the production over shovels, trucks' loading, travel (loaded and empty) and dumping times as well as their variabilities, shift's composition. The procedure models: the average duration of a complete truck cycle during stationary and non-stationary periods during the shift as a function of the number of trucks in the circuit, the composition of the average 100% operated shift, the simultaneous probabilities of operationally available shovels and trucks and the distribution of shovels and trucks according to production and cycle times. The procedure can be programmed for a small computer and requires only short computing time, which makes extensive studies of various parameter combinations practical to carry out. The development of the method is an extension of previous work by Morgan and Peterson {1}, who by simulation established the relation between the match factor (which is the ratio between the sum of one loading time pr truck in the circuit and the complete net cycle time) and the shovel's efficiency, and by Hayes, Splaine and Whitaker {2}, who established this relation from a large number of observations in the Nchanga pit in Zambia. The present method establishes this relation by a straightforward calculation procedure. Theoretical solutions to some situations are found in the litterature on the queuing theory. The papers {3},{4} and {5} give solutions to situations, which fall outside those encountered in the shovel-trucks circuit. These solutions together with simulations of in-termediary situations have been used to establish an interpolation procedure for the correction factors to be applied to the waiting time formula in order to cover the situations encountered in practice.

Jan 1, 1977

By Richard H. Sillitoe

Gold is an important byproduct and, locally, coproduct of porphyry copper deposits, and possesses potential for discovery as the major metal in porphyry deposits in either island arcs or Cordilleran belts. Tungsten and tin occur as major components of several potentially bulk-minable porphyry deposits, as well as being byproducts from the Climax porphyry molybdenum deposit. Whereas porphyry tin systems are localized in Cordilleran back-arc settings, their tungsten counterparts appear to span a range of geotectonic settings, from Cordilleran principal arcs through back-arcs to rifts. Although silver is a common byproduct of porphyry copper deposits, it tends to be preferentially concentrated with lead and zinc on the fringes of all types of porphyry deposits and/or as late-stage superpositions; the latter situation could be of some interest in the higher parts of porphyry systems. As byproducts, rhenium, uranium, platinoids and magnetite are exploited occasionally, and rutile and the rare metals, such as tantalum and niobium, have merited consideration.

Jan 1, 1982

By Walter N. Campbell

The Simon Silver Mine is located near Mina, Nevada, approximately 50 miles northwest of Tonopah and 40 miles east of Hawthorne. It is an old property which was last worked in 1932 and closed operations due to problems arising from low metal prices. Federal Resources Corp. acquired the property in 1963 and undertook a completely new development program since the old surface plant and underground workings were beyond feasible repair. The deposit of silver, lead, zinc and copper is predominantly replacement of limestone. In the summer of 1964, a new surface plant was constructed and on December 2, 1964, sinking of a three compartment vertical shaft started and was completed to a depth of 1088 feet in April of 1965. Following the sinking program an 850 foot crosscut was driven to intersect the mineral zone on the 1000 level after which drifting, crosscutting and raising totaling some additional 4280 feet has been completed along with considerable diamond drilling and long holing into the walls. To date, two ore bodies have been encountered and test stoping operations are underway together with development of the 800 level for servicing, ventilation and access. The present development of ore is directed towards substantiating sufficient reserves to warrant the construction of a flotation mill.

Jan 1, 1967