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By Robert D. Thomson

LISTED among strategic and critical minerals for stockpiling, mica is an important raw material in electronic and electrical equipment industries. Each year large quantities are consumed in the manufacture of television and radio tubes, radar sets, generators, transformers, electric toasters, and irons. Crude mica, which comes from the mine in crystals or in rough books of various shapes and sizes, is known as run-of-mine or book mica. Structural imperfections, impurities, rough edges, and defective laminae are removed by cobbing, rifting, and trimming to produce a marketable block mica or film mica.' After the rough books have been hand-cobbed to remove adhering rock, dirt, and crushed or defective mica they are rifted, or split. In the U.S. a straight-bladed knife is generally used for rifting. Products of this operation are classed broadly as untrimmed sheet mica and scrap. Sheet mica is any relatively flat piece sufficiently free from physical defects to be cut or stamped into specified shapes for industrial use. Sheet mica is classed as block, film, and splittings. Block mica is not less than 0.007 in. thick with a minimum usable area of 1 sq in. Film mica is split from the better qualities of block mica to groups of specified thickness ranging from 0.0012 to 0.004 in. Splittings are sheets with maximum thickness of 0.0012 in. and minimum usable area of 3/4 sq in. As virtually no mica splittings are produced in the U.S., they will not be discussed in this report. Small, low quality block is subdivided as punch and washer. Punch mica is thumb-trimmed material large enough to yield a circular disk of at least 11/2 in. diam for stained quality mica and 1 1/4 to 2 in. for clear quality. Poorer lots, often referred to as washer mica, are also included in the general term punch mica. Trimming, the next step in processing sheet mica, removes broken and ragged edges, loose scales, and other major imperfections. Trimming may be done with a knife, sickle, shears, or fingers. In the U.S. it is usually done with a 3-in. linoleum knife to obtain a bevel-edged cut. If ragged edges are broken off with the thumb and fingers, the mica is called thumb-trimmed. Full-trimmed is rifted mica trimmed on all sides with all cracks, reeves, and

Jan 1, 1957

By R. H. Hawkins

An investigation of the applicability of geophyiical methods to northern Ontario lignite deposits was undertaken early in 1930 by the Ontario Research Foundation at the request of the Ontario Department of Mines Pre-Cambrian ... Syenite, granite gnciss, etc. Igneous intrusivcs . Lamprophyre dikes cut the Abitibi River and lover formntions. (Permission of Ontario Dept. of Mines) and was financed in part from a research grant made by the latter organization. The work to date has included gravitational, magnetic and electrical resistivity studies. The gravitational and magnetic measure-

Jan 1, 1934

By R. V. Ramani, V. T. Burgos, J. A. McClay

This report was prepared by The Pennsylvania State University under USBM Contract No. G0111808. The contract was initiated under the Coal Mine Health and Safety Program. It was administered under the technical direction of DMSE, with Mr. G. R. Schottler acting as the technical project officer. Mr. J. A. Herlckes was the contract administrator for the Bureau of Mines. This report is a summary of the work recently completed as part of this contract during the period June, 1971 to September, 1974. Thls report was submitted by the authors on February 13, 1975.

Jan 1, 1975

By Hans Werner, Gunnar Almgren

An overview of the first Swedish Research mine at Mt Luossavaara, Kiruna, Sweden, is given, including its background, objective , program and expected results. The Research mine involves the spending of approximately US$ 32 Million over a five year period of which 30% are financed by the Swedish Government and remaining 70% by income from ore sales. The main objective will be to develop an underground mining system based on rock mechanical pre-investigations and blast hole stoping using big holes. The research program will include the following main components:

Jan 1, 1981

By A. H. Rhoades, D. T. King, R. J. Gray

Tests and methods of detecting oxidized coal are studied. Particular emphasis is placed on the microscopically discernible changes that accompany the chemical and physical changes affecting the technological properties of coal during oxidation. Coal samples were used from the low-volatile Pocahontas No. 3 Seam, McDowell County, W.V., and from the high-volatile Cedar Grove Seam, Mingo County, W.V. Tests conducted include: carbonization, free-swelling index, phase inversion, froth flotation, Gieseler plasticity, oxygen by difference and by neutron activation, volatile matter, pH of coal slurries, alkali solubility, and electrostatic charge.

Jan 1, 1977

CHEMICAL PRECIPITATION U.S. 4,067,789 - Removal of soluble manganous ions from a zinc sulfate solution formed by leaching roasted zinc sulfide ore with spent electrolyte. An excess of ammonium persulfate or like oxidant is added to the leach solution to convert manganese to insoluble, filterable gamma-manganese dioxide, and the precipitate separated from the solution by filtration. L. Harris and A. K. Hanson, Jr., assigned to Newmont Exploration Ltd. Jan. 10, 1978. 11 pp. (204-119). Filed: 5-3-76 (682.442). (Drawing on page 67.)

Jan 1, 1979

By P K. Seccombe

Sulfur isotope data are available for 18 low-sulfide orogenic gold systems hosted by the Lachlan Orogen in eastern New South Wales including those in the northern Hill End Trough (Hargraves, Hill End, Napoleon Reefs, Sofala-Wattle Flat, Stuart Town, Windeyer) and adjacent areas (Bodangora, Lucknow, Springfield), the Gilmore Fault Zone (Adelong, Sebastopol) and those in the Parkes district (Calarie, London-Victoria, Wyoming One-Myall United). Other systems for which data are available include Mongarlowe, Murrumbateman, McMahons Reef and Yalgogrin. The d34S data for many of the larger deposits including Adelong, Bodangora, Calarie, Hargraves, Hill End London-Victoria, Lucknow, Sebastopol, Sofala-Wattle Flat, Stuart Town and Windeyer (Group 1) are all very similar and have average d34S values close to zero per mil. Near 0ë d34S signatures imply that sulfur was sourced from a magmatic reservoir. Furthermore, the large Wyoming One deposit has d34S values significantly below 0ë with the data implying that sulfur was sourced from a fractionated magmatic reservoir. Group 2 deposits include those at Mongarlowe, Murrumbateman, Napoleon Reefs, McMahons Reef and Springfield . These deposits have average d34S values significantly above 0ë. With the exception of Springfield where sulfur was probably sourced from the Hill End Trough, Group 2 deposits are an order of magnitude smaller than Group 1 deposits. The results suggest that many of the larger orogenic-related gold deposits hosted by the Lachlan Orogen in eastern New South Wales derived sulfur and gold as a result of leaching and recycling from a magmatic reservoir. Alternatively, there may have been direct mantle-derived input; however, this is considered unlikely on geological grounds. An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation.

Jan 1, 2008

By Courtenay de Kalb

So far as the writer is aware, no detailed investigation into the behavior of ores or rocks when subjected to crushing under digerent conditions has yet been made. He cannot himself claim to have carried such investigations very far; but the results of some studies which he has been pursuing for a couple of years, as opportunity permitted, seem sufficiently interesting and suggestive to warrant a brief statement of them, preliminary to nildertaking a more exhaustive course of experiments. Criticisms and suggestions are earnestly invited, with a view to making these studies as instructive as possible. As a part of the examination of ores, preparatory to concentration, my students have been required to take lots of 100 pounds each, and crush them. These are then screened through a series of hand-sieves, 15 inches in diameter. The products remaining on each screen are weighed, and from the results is plotted a curve, called the sizing-curve. It was assumed at first that different ores would yield, under like conditions of crushing, curves showing unlike characteristics as to the quantities remaining on the various sieves. After a number of such curves had been obtained it became evident, as often happens, that this a priori reasoning was faulty. The curves presented practically identical characteristics, and only by varying the crushing-conditions were variations in the curves obtainable. Another peculiarity, which was as little anticipated, appeared by inspection of these results, namely, that successive crushing to smaller sizes did not greatly augment the quantities of the filler products, so long as the machines employed applied the crushing-force to the ore-particles in the same manner. In these experiments the fines were not eliminated between successive crushings. I believe it will be readily admitted that rock can be crushed

Jan 1, 1899

By Holbrook G. Botset

Experiments performed in this laboratory on the flow of gas-liquid mixtures through unconsolidated sands have been described and discussed in an earlier paper.4 In these earlier experiments a definite relationship was found between the liquid saturation of a sand and its permeability to the liquid or gas phase. Furthermore, the nature of the sand, the permeabilities of the sands over a range between 17 and 260 darcies, appeared to have a negligible effect upon the saturation-permeability relation. Likewise experiments indicated that the viscosity of the liquid, or the substitution of an immiscible liquid for the gas phase, had only a very slight effect on the permeability-saturation relation. This conclusion has since been definitely confirmed by the experiments of M. C. Leverett3 on the flow of oil and water through unconsolidated sands. The sands used by Leverett were of very low permeability (between 6.8 and 1.04 darcies), while the oil-water viscosity ratio varied between 90 and 0.057. Since the sands that constitute petroleum reservoirs are largely consolidated, although of widely varying degrees of consolidation, it was desirable to extend the experiments on gas-liquid mixtures to consolidated sand. In the study of consolidated sands additional variables are involved, but for the initial experiments the only new variable to be introduced was the cementing material. Experimental Procedure A comparatively limited assortment of sandstones was available for the selection of the experimental material. Of these, the Nichols buff sandstone appeared to be most suitable. It gives no visual evidence of any bedding planes, it has a firm structure, permitting easy handling, and its permeability is about 0.5 darcy. Its only disadvantage is the fact that it contains an appreciable amount of ferric oxide which, for experiments involving water, will hydrate and reduce the permeability. This

Jan 1, 1940

By Holbrook G. Botset

Experiments performed in this laboratory on the flow of gas-liquid mixtures through unconsolidated sands have been described and discussed in an earlier paper.4 In these earlier experiments a definite relationship was found between the liquid saturation of a sand and its permeability to the liquid or gas phase. Furthermore, the nature of the sand, the permeabilities of the sands over a range between 17 and 260 darcies, appeared to have a negligible effect upon the saturation-permeability relation. Likewise experiments indicated that the viscosity of the liquid, or the substitution of an immiscible liquid for the gas phase, had only a very slight effect on the permeability-saturation relation. This conclusion has since been definitely confirmed by the experiments of M. C. Leverett3 on the flow of oil and water through unconsolidated sands. The sands used by Leverett were of very low permeability (between 6.8 and 1.04 darcies), while the oil-water viscosity ratio varied between 90 and 0.057. Since the sands that constitute petroleum reservoirs are largely consolidated, although of widely varying degrees of consolidation, it was desirable to extend the experiments on gas-liquid mixtures to consolidated sand. In the study of consolidated sands additional variables are involved, but for the initial experiments the only new variable to be introduced was the cementing material. Experimental Procedure A comparatively limited assortment of sandstones was available for the selection of the experimental material. Of these, the Nichols buff sandstone appeared to be most suitable. It gives no visual evidence of any bedding planes, it has a firm structure, permitting easy handling, and its permeability is about 0.5 darcy. Its only disadvantage is the fact that it contains an appreciable amount of ferric oxide which, for experiments involving water, will hydrate and reduce the permeability. This

Jan 1, 1940

By Oliver C. Ralston

A definite field of usefulness has developed for the brine-leaching processes of removing lead from ores and other products, so this paper reviews the developments, both in practice and in research, that have taken place since the publication of two earlier papers on this subject.' Several commercial plants are now in operation and a number of semi-commercial test plants have been operated in an effort to determine the limitations of brine leaching and to compare the results with other forms of treatment. Briefly, the brine leaching of lead depends on the fact that lead chloride and lead sulfate are soluble in saturated sodium-chloride solutions. In dilute brines, the solubility of the lead salt is less than in pure water; but in the strong brines, the solubility is greater than in pure water due, supposedly, to the formation of a double sodium-lead chloride. This is seen in Fig. 1, which shows the results of M. Demassieux2 on the solubility of lead chloride in sodium-chloride solutions of different concentrations and at three temperatures. The curves are not those given by Demassieux but are calculated from his results in order to convert them to units common in metallurgical work. The right-hand branch of each curve shows that if the solution is kept in contact with solid sodium chloride the lead chloride can be almost entirely displaced from solution, but that brines which are saturated with sodium chloride, but are not in contact with the solid salt, have a very high solubility for lead chloride. The diagram gives the analysis of the solution in grams of lead per liter, although the substance actually in solution is lead chloride. At 100" C. as much as 100 gm. of lead per liter can be obtained; but at lower tem-

Jan 1, 1924

By Charles W. Finkl

Coastal and marine aggregate (sand and gravel) mining on US continental shelves is a major activity that supplies industrial and commercial demands for these raw materials. Because Pacific, Atlantic, and Gulf coasts contain significant shoreline segments that are in a critical state of erosion, sand-gravel and mixed sediment (sands with a high percentage of silt-plus-clay) deposits are mined in efforts to mitigate land loss. Sandy shores, and to a limited extent gravel coasts, erode by shoreline retreat due to relative sea-level rise and sediment starvation in littoral drift patterns. Mostly gravel shores occur in high latitudes and shorelines often advance seaward due to isostatic rebound following the LGM (Last Glacial Maximum 18,000 YBP). Sandy and fine-grained shores occurring in middle and low latitudes of the conterminous United States are problem coasts that require beach renourishment and reconstruction of barrier islands. Muddy coasts of the Mississippi Delta region are especially problematic because some islands are literally disappearing in response to subsidence. Replacement of sediments lost to the littoral drift system or returned to sedimentary sinks by drowning is thus a critical endeavor that involves annually dredging hundreds of millions of cubic meters of sediments from the continental shelf area. The types of shelf sediments that are amenable to marine mining techniques, their occurrence, and means of detection are highlighted in this paper. Minable Sediments on the Continental Shelf Wide varieties of sedimentary bodies that occur on the continental shelf are amenable to mining for beach renourishment and barrier island (beach-dune-marsh system) reconstruction. Relict terrestrial deposits make up a small but important proportion of shelf sediments that were deposited on coastal plains, now drowned by the post-glacial sea-level rise to form continental shelves. Examples of these deposits include drowned moraines, outwash plains, eskers, and kettles along the northern Atlantic coast in addition to siliciclastics along middle latitude Pacific, Atlantic, and Gulf coasts. Special cases include carbonate sequences along tropical costs of Florida and the muddy coasts of deltas, mostly in Louisiana. Specific types of minable deposits thus include shoreface attached sand sheets, dunes, and ridges in addition to offshore inter-reefal sands (between coral reefs in the Florida Reef Tract), bank deposits, and sand waves (ridges) that extend into US territorial waters. Nearshore siliciclastics include ebb-tidal deltas and bars. Drowned river valleys occur along all three US coasts and can be exploited for their fluvial sands if the covering mud sheets are thin.

Jan 1, 2005

By José Aparecido Pereira

The customer asked for a 5” angle roll pass design. For this angle, in a normal pass design, a 160mm billet is required. For the production of this billet would be necessary a new 160mm mold for the Continuous Casting Machine, which produces 150mm billets, increasing the complexity of the melt shop. The rolling mill is designed to produce angles up to 4", technical information recommended the replacement of a gearbox because this process would generate torque up to 15% above nominal. An innovative roll pass design was done using 150mm billet using the Aparecido-W Method that cause a higher spread in the first three passes, thus reaching the desired leg length. As this method is not used in any other known rolling mill, a FES (Finite Element Simulation) and FEM (Finite Element Modeling) was done by a Research & Development center which demonstrated that the design was viable. From July, 2018, three tests were done being produced around 700 t of good quality products without difficulty of set up. The rolling mill team carried out this project at a very low cost and in record time of five months. The average torque of the motors was 36%, without overload for the equipment.

Oct 1, 2019

By B. T. Brady, Wilson Blake

One of the more widespread experimental procedures currently used in experimental rock and soil mechanics work involves the use of a short circular cylinder loaded axially between nonrigid end plates1-5 (Fig. 1a and b). It is widely recognized 6 that a number of undesirable features are connected with this test procedure ; the most noteworthy of these is the development of a nonuniform stress distribution throughout the specimen resulting from an elastic mismatch between the specimen and the end plates. In general, the end plates and specimen possess different elastic properties. As a result of this elastic mismatch, the uniform distribution of stress and strain applied to the end plates is altered at the interface, the degree of which is dependent not only on the magnitude of the elastic mismatch but also on whether there is any relative movement between the specimen and the end plate. A convenient measure of the elastic mismatch is the ratio (vp/Ep) / (v8/E3), where v and E are Poisson's ratio and Young's modulus and the subscripts s and p refer to the specimen and end plate, respectively. For most tests that use steel end plates, (v8/E8) is greater than (vp/Ep) ; i.e., the specimen tends to expand or contract laterally more than the end plates. Because of the frictional restraint offered by the end plates, shearing stresses are developed at the interface that, for compression testing, act inward producing a "clamping" effect (Fig. lc). The effect of this "clamping" is to produce a state of triaxial stress throughout the specimen. Fig. Id illustrates the coordinate system and stress nomenclature used. In a homogeneous, elastic two-dimensional plate loaded uniaxially at the edge, a uniform stress distribution is obtained at a distance from the

Jan 1, 1972

By H. K. Taylor

"As a mining engineer with over 50 years' experience in underground operations and mine planning, I appreciate the enormous amount of work that went into the paper; and despite an imbalance between geology-geostatistics and the profit-earning actions of mining and treating (possibly because of sponsorship only by the Geological Society), I agree with probably 85% of the detail. But the paper fails to build on its own sound foundations and, like many predecessors, drifts instead into idealism. Even the starting quotation from the US SEC Form S-18 is little more than a platitude. An American consultant (l) has observed: ""an evaluation that accepts ore reserves as guaranteed because their definition meets SEC criteria is an invitation to disaster"".Idealism enshrines at least two myths that have bedevilled definition over the last several decades. The first is that the only purpose of 'ore reserves' is to promise successful new mines, and that 'ore reserves' must necessarily follow feasibility studies.In practice, the identification of what is likely to be mined if a mine is started must often be done on broad economics long before the feasibility study, and there may be only little later flexibility. The second myth is that there can be a perfect deposit with a simple exploration-appraisal-development sequence (the paper's Figure 2) against which all inferiors can be matched. If the field of definition be regarded as a pyramid, the quest for perfection and the imposition of rigid promises upon the terms 'ore' and 'reserves' drive definitions so near the apex (or the bottom right corner of Figure I, if preferred) that little is left to define. The process is reminiscent of the legendary super-medicine, the one so powerful that the patient had to be in perfect health to take it!"

Jan 1, 1992

By C A. Thomas, J C. Eardley, G V. Franks

As valuable minerals are more disseminated, there is a need for low-grade ores to be finely ground to liberate these valuable minerals. However, this has a negative impact on ultrafine mineral recovery as conventional separations processes such as gravity separation are no longer effective. Instead, fine particles can be recovered by froth flotation. Various novel reagents have been introduced along with development in flotation cell designs such as Jameson cell and column cells to improve performance and recovery of ultrafine valuables. This study, on the other hand, aims to recover the fines using a combination of commercially available reagents to selectively flocculate the valuables, so they are large enough to be recovered via conventional froth flotation cells. This technique if successful can be implemented in many existing industries that continue to use mechanical flotation cells that require particle sizes to be 50–150 μm for effective recovery (Ralston, 1992; Batterham and Moodie, 2005). Selective flocculation is not widely implemented in industries mainly due to the inability to float the aggregates ones they are formed and gangue entrapment within the aggregates which consequently affects the grade of concentrate recovered (Forbes, 2011). The major goal is to aggregate the minerals and float the aggregates unlike the more common industrial technique such as reverse flotation of hematite, where the aggregates itself is not recovered via flotation.

Aug 24, 2022

By J. P. Newell, R. W Storey

This paper describes the development of a continuous float-and-sink process to produce coal low enough in ash content to be suitable for production of electrodes. The cleaned coal had a combined iron and silicon content of 0.239 pct. The efficiency of recovery was 84 pct. AUGER mining is a form of continuous mining in that it completely replaces with a one-cycle operation the older conventional cut, drill, shoot, and load method of mining. Relatively new, having been utilized only within the last few years, it is even more continuous in its production of coal than some other forms of continuous mining being practiced in this country today. There are at present coal recovery augers in operation working along strip-mined highwalls from the outside. This paper will deal with underground mining, specifically with an experiment which was carried out with a machine developed for this purpose. The location of this experiment was in Pike County in eastern Kentucky in the upper split of the No. 3 Elkhorn seam. Coal height was 32 in. and roof conditions were very bad, with 10 in. of soft draw slate coming down with every cut in conventional mining. The coal is high-grade bituminous, excellent for metallurgical coking. Underground auger mining fills a need for a method of recovering coal in thin seams, under bad roof conditions or in split seams. In many cases it might save money by eliminating the handling of excessive quantities of rock. In fact, if the coal can be recovered, as in the case of this experiment, without mixing with any impurities outside the seam, it may be possible to load or ship a satisfactory product with little or no preparation or cleaning expense. The need for some economical method for recovering coal from thin seams is becoming more pressing as the more profitably mined seams are exhausted. Basic Principles of Auger Mining The principles of the auger, or screw conveyor, have long been used in drilling small diameter holes for blasting and other purposes. These same fundamentals apply to large diameter drilling. However, considerably more attention was given to designing a cutting head that would not pulverize the coal and consequently would require a minimum of power to drive it. The head, as designed for this job, consists of a cylinder 25 in. in diam with steel bits set in 3 positions spaced around the forward rim. A conventional coal drill bit, at the head of a bursting cone in the center of the cylinder, trails the forward rim slightly, see Fig. 1. The bursting cone and auger connections are anchored to the cutter barrel by 3 arms spaced at 120". This head design proved to be very satis-

Jan 1, 1953

By J. P. Newell, R. W. Storey

This paper describes the development of a continuous float-and-sink process to produce coal low enough in ash content to be suitable for production of electrodes. The cleaned coal had a combined iron and silicon content of 0.239 pct. The efficiency of recovery was 84 pct. AUGER mining is a form of continuous mining in that it completely replaces with a one-cycle operation the older conventional cut, drill, shoot, and load method of mining. Relatively new, having been utilized only within the last few years, it is even more continuous in its production of coal than some other forms of continuous mining being practiced in this country today. There are at present coal recovery augers in operation working along strip-mined highwalls from the outside. This paper will deal with underground mining, specifically with an experiment which was carried out with a machine developed for this purpose. The location of this experiment was in Pike County in eastern Kentucky in the upper split of the No. 3 Elkhorn seam. Coal height was 32 in. and roof conditions were very bad, with 10 in. of soft draw slate coming down with every cut in conventional mining. The coal is high-grade bituminous, excellent for metallurgical coking. Underground auger mining fills a need for a method of recovering coal in thin seams, under bad roof conditions or in split seams. In many cases it might save money by eliminating the handling of excessive quantities of rock. In fact, if the coal can be recovered, as in the case of this experiment, without mixing with any impurities outside the seam, it may be possible to load or ship a satisfactory product with little or no preparation or cleaning expense. The need for some economical method for recovering coal from thin seams is becoming more pressing as the more profitably mined seams are exhausted. Basic Principles of Auger Mining The principles of the auger, or screw conveyor, have long been used in drilling small diameter holes for blasting and other purposes. These same fundamentals apply to large diameter drilling. However, considerably more attention was given to designing a cutting head that would not pulverize the coal and consequently would require a minimum of power to drive it. The head, as designed for this job, consists of a cylinder 25 in. in diam with steel bits set in 3 positions spaced around the forward rim. A conventional coal drill bit, at the head of a bursting cone in the center of the cylinder, trails the forward rim slightly, see Fig. 1. The bursting cone and auger connections are anchored to the cutter barrel by 3 arms spaced at 120". This head design proved to be very satis-

Jan 1, 1953

By H. G. Naidu, Raj K. Singhai, S. P. Singh

Introduction The economics of open-pit mining de¬pend heavily on the cost of overburden removal. It is necessary to maintain a pit slope as steep as possible. Steeper slopes lower the volume of overburden to be handled, thus lower the cost of material handling. There is, however, a limit to which the pit slopes in any given situation can be steepened. Steepening beyond this point could jeopardize the pitwall stability, ren¬dering it unsafe. It might even result in curtailment of pit production. Several factors affect the stability of pitwalls. Some are natural, such as topogra¬phy and regional and local geology. Others are man made, such as pit geometry, ground water management, equipment, drilling and blasting, and the type of excavation. Greater control can be exercised in the latter with proper planning and by following a set of procedures based on field experiences. This article deals with one such control¬lable factor - drilling and blasting. The eco¬nomics of steeper pit slopes is discussed, followed by a record of drilling and blasting practices in western Canadian surface mines. Microprocessor-based instrumentation and its application to blasthole drilling is also described. Economics of steeper pit slopes The economics of steeper pit slopes is not a new issue. It should be reinforced, though, that pit slopes must be as steep as possible to ensure maximum recovery of minerals at a minimum stripping ratio. For operating, economic, and safety reasons, different pit slopes are maintained during different phases of pit development. Figure 1 shows a hypothetical ore body, cylindrical shape, with a radius of 225 m (740 ft) and located 30 m (100 ft) beneath the surface. The mining operations are planned at two different slope angles of 25° and 40°. The ultimate pit depth was established on the basis of the criterion of "break even pit." The variation of the stripping ratio at differ¬ent mine levels is shown in Fig. 2. Table I summarizes the eco¬nomics involved for two pit slopes, 25° and 40°. Figure 3 presents the profits during the life of the mine for these two pit slopes. This study shows that steeper slopes for an open pit operation are far more advanta¬geous than flatter slopes. But, what is the optimum slope at any one point in the life of a moving pit? And how can the stable slopes be maintained over the short- and long-term? Canadian blasting practices A survey of blasting practices in western Canadian surface coal mines revealed that, at most mines, hard consolidated overburden is drilled and blasted. Blastholes are drilled with rotary crawler-¬mounted rigs using bits in the 220 to 310 mm-diam (8.5 to 12 in.-diam) range. Typi¬cal hole depths are 12 to 18 m (40 to 60 ft) to accommodate the pit bench heights of 12 to 15 m (40 to 50 ft). For a 12-m (40-ft) high bench, a typical hole will be 14.6 m (48 ft) deep including the subgrade drilling of 2.6 m (8.5 ft). Half of this hole will typically be a powder column and half will be stemmed. For a 15-m (50-ft) bench, a typical hole will be up to 18 m (66 ft) deep including the subgrade drilling. Drilling patterns are usually 9 m (29.5 ft) spacing and 9 m (29.5 ft) burden. De¬pending on local geology and hole diameter selected, drilling patterns of 8 x 8 m, 8.5 x 7.3 m, and staggered patterns of 11 x 5.5 m

Jan 1, 1989

By G. A. Burrell, G. W. Jones

As a part of the work which the Bureau of Mines carried out in 1916 and 1917 pertaining to the compressibility of natural gas , it became necessary to de termine the pressure -volume relations of several of the paraffin hydrocarbons of which natural gas is composed . Results of the work on natural gas were published in Technical papers 131 (Burrell , G.A. , and Robertson , I.W. , The compressibility of natural gas to high pressure , Bureau of Mines Tech . Paper 131 , 1916 , 12 pp. ) and 158 (Burrell , G.A. , and Robertson , I.W. , Compressibility of natural gas and its constituents with analyses of natural gas from 31 cities in the United States , Bureau of Mines Tech . Paper 158 , 1917 , 12 pp . ) , and curves were given of pressure volume deviations of the gases , methane , ethane , propane , and carbon dioxide at high pressures . original experimental data on these gases , however , were not published . The In view of many inquiries which the Bureau has received with reference to the original data on the se gases , and the importance of the pressure -volume relations from which it is possible to calculate the deviation of any natural gas who se composition is known , the original data as determined by Burrell and Robertson are pre sented below .

Aug 1, 1921