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By George Holderer

THE Enos Coal Mining Co. owns an area underlaid by coal, of approximately 5000 acres. This property is situated 9 miles from Oakland City, in the southern part of Indiana. The mine has been in operation 5 years, during which time some 350 acres have been stripped, and stripping is now proceeding at the rate of about 100 acres per year. TOPOGRAPHY These coal lands consist of gently-rolling agricultural fields which in times past were farmed for wheat. There is no timber and but very little brush to contend with. The overburden varies from 10 to 60 ft. in depth. A thin bed of slate overlies the coal, followed by limestone, which varies from 3 to 15 ft.; this, in turn, is followed by sandstone, slate, shale and clay. There is a very wide variation in the thickness of all of these beds. The coal has an average thickness of 5 ft., and is a good grade of high-volatile bituminous. MINING OPERATIONS Briefly, the sequence of operations is: drilling holes with churn drills; blasting with L. 0. X.; stripping overburden to uncover coal and, finally taking up the coal, which is sent to a tipple for cleaning and sizing. In order to show what is being accomplished by the use of L. 0. X., results will be shown from May to November, 1926, inclusive, and compared to the same period in 1925. No work was done in No. 1 pit during May, 1925, hence all figures are unduly low for this month. Drilling There are in use 8 No. 44 Loomis clipper drills, caterpillar traction with gasoline engines; recently these were equipped with extra heavy stems, measuring 26 ft. overall and 4 ½ in. dia., to increase the speed of drilling. In addition, 4 No. 1 ½ G non-traction drills have been acquired, and are used to save the time of the larger drills when recleaning is necessary.

Jan 1, 1927

By R. S. French, W. R. Hibbard

FOR tensile deformation, if the stress value is defined by the ratio of the load to the actual area, and the strain value by the natural logarithm of the ratio of the immediate length to the original gauge length of the sample, the resulting data when plotted with logarithmic coordinates result in a linear relation extending from the initial plastic yielding through the maximum load and in some cases, such as for copper and brass, to fracture'.'. The normal form of the curve, therefore, is a parabola, satisfying the following equation: S = Kem [1] where S and e represent true stress and strain, K is a constant, and m is a coefficient evaluating the slope of the curve and designated therefore as the strain hardening coefficient. Hollomon² has reported that this coefficient is a function of the yield strength as effected by grain size in alpha brass, and the carbon content and heat treatment of steel. In the case of steel, the value of the proportionality constant is a function of the carbon content. It was further shown in that paper2 and also by Low and Prater -hat, if the power relation between stress and strain remains valid, the strain hardening coefficient m, is equal to the strain at the maximum or necking load. Thus, an added significant feature is that this coefficient may serve as an index of the capacity for deformation prior to localized necking and subsequent fracture, e.g. drawability. The constant K may be defined as the strength of the material at unit strain. Together the K constant and the m coefficient describe an alloy's physical properties as far as strength and ductility are concerned for a particular yield strength and, in the case of steel, a particular carbon content. The work of Lacy and Gensamer4 gives data on alloy ferrites from which very close estimates of the strength of ternary and higher iron alloys have been made. It was the purpose of this investigation to develop similar data for commercial copper base alpha solid solution alloys. Preparation of Specimens Commercial binary alloys of copper with aluminum, beryllium, cadmium, nickel, silicon, tin and zinc were investigated. In addition, laboratory castings were made of a 1 pct silicon alloy and a 2 pct aluminum alloy to complete these particular series. Two types of copper were used, oxygen free OFHC and tough pitch FC. The analyses of these alloys are shown in table I. Standard ASTM sheet tensile specimens were used to determine the usual physical properties of each alloy and to obtain stress data at strains from 0.0005 to 0.01. Small (1/4 in. x 3 in. x tk.) samples

Jan 1, 1951

By J. B. Dilworth

I. INTRODUCTION. THE purpose of this paper is, first, to give a general account of a little-known coal-district of SE. Kentucky, its topography, drainage, and mineral resources, for those who may be interested in its economic development; and secondly, so to describe the geologic features of the field as to furnish available data to those engaged on the broader problems of Appalachian coal-geology. The data from which the major portion of the paper has been prepared were obtained, in professional investigations for private individuals and corporations, within the last decade. The area under review, defined by a heavy line on the map, Fig. 1, lies between Little Black and Pine mountains, and extends from Harlan on the west to a N-S. line crossing Looney creek 3 miles from its mouth on the east. This area can be divided for description into a Western and an Eastern Division. The character of investigation pursued in the Western Division, below Fugitt creek, was very different from that employed in the Eastern. In the former, only a rapid reconnoissance was attempted : such coal-openings as had been made by prospectors or by citizens in obtaining fuel were visited, their sections measured and, in many instances, sampled. Elevations were obtained barometrically and approximate locations sketched on a small-scale map. In a field where the coals are so well opened, the reconnoissance method gives a very fair idea of the number and character of the principal beds and their relative positions, and affords much evidence for their correlation over wide areas, without permitting the more refined work required to discover all the seams of the measures and to determine the details of geologic structure.

Feb 1, 1912

By Lewis M. Haupt

Discussion of the paper of Mr. Granger, presented at the New Haven meeting, February, 1909, and published in Bulletin No. 25, January, 1909, pp. 1 to 37. LEWIS M. HAUPT, Philadelphia, Pa. (communication to the 'Secretary*) :-This subject is of great interest to me, and I am noting progress and conditions as a test of my own judgment in advocating what I believed to be the nearer, cheaper and better route at Nicaraugua. I have as yet learned nothing to change my views, or modify the convictions which led me to retire from the Commission; but that is not the issue to-day. Our government is committed to the Panama route, and is bound to " make good." There are many points presented by Mr. Granger which seem to me worthy of very serious consideration. But when all is said there remains the risk of seismic convulsions, to which either form of canal is subject; and I fail to see that the risk would be reduced by making the cut deeper, with greater surcharge of lateral pressure upon the side walls. A severe earthquake in the rainy season would doubtless close either type of canal. One of the arguments in favor of the Panama route most insisted on, was that Nature had provided good natural harbors at its termini. Recalling this, it has interested me to note that the cost of creating the necessary harbors is estimated at $27,000,000. Mr. Granger refers to this, and quotes the Board's report to the effect that, on the Atlantic side, it will require 5 miles of jetties to " make an excellent harbor out of a wretched one." The vulnerable part of Mr. Granger's paper seems to me to lie in the admission that the plant required is still in the embryonic stage, and that at least one machine should be built and tested " to demonstrate that the theoretical strains and forces work out in practice exactly as expected." This, how-

Jul 1, 1909

By Peter F. Weiss, H. Riedler, I. M. Moschitz, A. Olsacher, G. B. Fettweis

CONTENTS : 1. NATURAL FACTORS Geology Petrography and Rock Mechanics Mining System Orientation Rock Bolting System Movability of the Ore Grain Size Distribution Permeability Contamination Value of the Ore Hydrology 2. TECHNICAL FACTORS Height Length and Width Drawing Areas Time Factor Water Predrainage Surf ace Draining Underground Draining Advancing Slitting Sequence of Development Initiation of the Caving Process Drawing Rate 3. SPECIFIC MINING RULES Advancing Development Drawing Rate Interruption of Drawing (Water Distribution, Concentration of Mud, Dilution) Blasting 4. CONTROLS Safety Controls for Operation Computerised Drawing Control Structure of the Computer Program Hard Ware Print Outs 5. SUMMARY 6. ACKNOWLEDGEMENTS 7. REFERENCES Under the heading "Development System for Block Caving under Severe Conditions" a report was presented to this congress regarding a magnesite mine using block caving for several years. On account of the experience gathered in this mine and the results of far reaching exchange of experience, as well as under due consideration of the international literature it is tried to systematically list and briefly discuss the following factors that have to be considered for development and draw control of block caving. 1. NATURAL FACTORS It is self evident that this implies all those circumstances given by nature and which cannot be altered by man. Any mining activity is subjected to those and the success is depending on how far technology can be adjusted to those conditions. Geology Deposits of considerable thickness with rather steeply inclined flat-faced contacts are optimal for block caving. While vertically there is practically no limitation- for the deposit, theoretically, the smallest horizontal basis apt for operating block caving is the minimum area necessary to get the rock to completely cave in by undercutting and drawing. Practically, only consider- ably bigger main dimensions are found, mainly because of technical factors (machinery and equipment) and economic factors (cost ratio of exploration vs. development, danger of contamination).

Jan 1, 1981

By A. Faber du Faur

At the request of some of the members to whom I have shown a collection of specimens of sulphur from southern Utah, I herewith present some notes relating to the mines and the method of extracting the sulphur from the ore. In preparing these notes I have borrowed from a paper by Prof. G. vom Rath, published in the Neue Jahrbuch fur Mineralogie, vol. i., 1884, dated Salt Lake City, December, 1883. The Cove Creek sulphur-bed, now worked by the Dickert & Myers Sulphur Co., is in Beaver county, near the line of Millard county. It was first discovered by Mr. Ferdinand Dickert in 1869 ; other beds in the same neighborhood mere found by him from 1872 to 1875, making altogether 22 mining claims. The want of railroad communication prevented the early development of these claims; but in 1883 Mr. F. Dickert put up a plant for extracting sulphur from the ore, which proved a perfect success; and in 1885 the property was transferred to a stock company organized under the laws of Utah. The railroad station nearest to the Cove Creek mines is Black Rock, on the Utah Central Railroad, 204 miles south from Salt Lake City, and about 26 miles west of the mines. The road from Black Rock to the mines passes through desert and lava-bed, to the road leading from Kanosh to Beaver, which latter road it joins near Cove Creek fort, a square stone enclosure, erected by the Mormons for protection against the Indians. Cove Creek fort is situated near the eastern limit of a plain bounded on the west at a distance of about 12 miles by the Mineral range, and on the east by the Sulphur mountains, forming part of the Tushar mountains. To the northwest the plain runs into the Sandy desert. About 3 miles from the fort rises a solitary hill about 1000 feet above the

Jan 1, 1888

By B. C. Allen, R. I. Jaffee, D. J. Maykuth

Wrought unalloyed iodide chromium, containing 39 to 95 ppm total interstitials, has a tensile transition temperature of —15°C. Re crystallizing at 1100°C causes the transition to rise to 90° to 390°C, depending on the purity, grain size, and cooling rate after annealing. At about the 100 ppm level, individual additions of carbon or nitrogen raise the transition of both wrought and recrystallized chromium by as much as 210°C, while oxygen has a lesser effect, and sulfur essentially no effect. High transitzon temperatures are exhibited by recrystallized and quenched chromium, and appear to be associated with the amount and distribution of inter-stitials. The latter cm be either in solid solution or in the form of precipitates within the grains or at the grain boundaries. Quench hardening in chromium is primarily caused by nitrogen. The hardness, re crystallization, and grain growth behavior of chromium with and without impurity additions were also investigated. BECAUSE of low density and good oxidation resistance, chromium-base alloys are potentially useful for elevated-temperature applications. The main deterrent to the use of chromium in structural applications is its lack of low-temperature ductility, especially in the recrystallized condition. Accumulated evidence has shown that the ductility of chromium is sensitive to purity. The objective of the current study was to determine the effect of common nonmetallic impurities on the tensile properties of high purity chromium. The separate effects of carbon, oxygen, nitrogen, and sulfur additions on the ductile-to-brittle transition were evaluated for wrought and recrystallized material after various cooling rates. EXPERIMENTAL WORK Unalloyed iodide chromium and seven impurity alloys containing individual carbon, oxygen, nitrogen, or sulfur additions were prepared in rod form. Nominal additions were 100 ppm C, N, and S, and 100 to 500 ppm O. The additions were about a factor of ten greater than the residual amounts in the unalloyed material. The seven alloys and two unalloyed compositions were consolidated into four-pound ingots by conventional consumable arc melting of pressed and welded electrodes under 0.5 atm of helium. Impurity additions were made to the electrode in the form of crushed master alloys. The impurity alloys were double melted to improve homogeneity. The ingots were broken down by extrusion at 1200°C in evacuated steel cans at a reduction ratio of 18 :1, using a molten borosilicate glass lubricant and a ram speed of 80 in. per min. The cans were removed by acid pickling, and the chromium core ground free of surface undulations. Cut lengths were sheathed in mild steel and swaged at 900°C until the core was about 1/4 in. in diam. Two additional 15-lb ingots of unalloyed chromium (ICr-3, ICr-4) were prepared by arc melting and broken down by 26:1 extrusion at 1250°C without a steel can. The billet was lubricated and protected from contamination by a coating of molten glass. The resulting extrusion was then sheath swaged as described previously. Representative chemical analyses of each of the materials, after removal of a 10-mil surface layer

Jan 1, 1963

By L. W. Rowe, S. S. Cole

A laboratory bench scale unit is described whereby finely divided chlorinatable residues are held for a short period by a restraining bed of a coarse-grained ore of comparable composition to permit 'flash' chlorination of the high surface area powder. Results on chlorinating a finely divicled titania residue using rutile sand as the restraining bed showed that 92 pct utilization of chlorine and equivalent conversion of TiO2 in the fine residue to Ticl4 were attained with a dust loss of about 3 to 4 pct and an equivalent amount of ch1om'nation of the titanium values in the restraining bed. THE chlorination of coarse grain titaniferous ores; i.e., 20 to 150 mesh sizes in fluidized beds is practiced commercially.' One effective method found for the chlorination of a very fine material in a fluidized bed is the "flash" chlorination technique.' By charging the finely divided charge material into the bottom of an established bed of optimum sized rutile, ilmenite, titanium beneficiate, or even an inert material such as sand, chlorination occurs before the dust can pass through the bed. The major requirement is that the bed be less chlorinatable than the charge. In this process, the charge consists of a finely divided (between one and 50 µ) titania residue obtained from a titaniferous ore beneficiation process. The charge is blended with a 20 pct wt basis addition of 20 to 60 .mesh size low volatile coke (to minimize hydrogen chloride formation) and fed to the bottom of the bed with the chlorine gas stream. A ball check device is used to permit entry of the charge and prevent the bed from discharging when the gas flow is cut off. The coarse pieces of coke serve to reduce compaction in the charge mixture and permit sat]sfactory feeding without plugging or jamming. Feed and gas rates are adjusted to provide for complete reaction of the charge material, with 90 to 95 pct chlorine efficiency, and maintain a gas flow in the range of 0.4 to 0.6 ft per sec. The equipment consisted of a 3-in. diam mullite tube enclosed in an electric furnace. A bow-shaped 3/4 in. Saran tube fed the charge from an air lock type of feed hopper (to prevent the escape of the chlorine gas into feed bin) to the bottom of the bed. The charge rate was set by providing a small awiliary hopper below the feed hopper where a weighed portion of the charge could be discharged in a specified time interval into the main chlorine stream. The chlorine rate was measured by a flowmeter. The reaction tube was filled with a bed of Australian rutile ore, this restraining bed serving two purposes; i.e., it rapidly heated the fresh charge to the reaction temperature and retained the particles in the bed for a sufficient time for selective chlorination to proceed. Due to the fine particle size and open, porous structure of the charge, this reaction was almost instantaneous, and, with proper adjustment of feed and gas rates, very little fine material passed through the bed without reacting and only a very small fraction of the restraining bed was chlorinated. A typical reaction procedure was to fill the bench scale fluidized bed unit with a 1 kg charge of Australian rutile ore analyzing 95.5 pct TiO2, 0.6 pct FeO, a beach sand deposit 40 to 150 mesh size. When fluidized, the bed was approximately 2 ft in depth. The ore was fluidized with a 1:l mixture of chlorine and nitrogen at a rate of 0.40 ft per sec. This gas mixture was necessitated by the comparatively short bed depth; a larger unit, with a deeper bed, could be fluidized with chlorine alone without significant free chlorine loss. Feed rate was 8.0 g per min of beneficiated ilmenite analyzing 85.0 pct TiO2, 2.6 pct FeO plus 1.6 g per min of 20 to 60 mesh petroleum coke. The

Jan 1, 1962

By E. E. Donath

Application of the coal hydrogenation process for the production of chemicals is described. It has been estimated that a plant to produce 31,090 bbl per day of chemicals and fuels would cost $326,- 000,000. These products would consist of 6.1 pct by weight tar acids, 50.8 pct aroma tics, 16.4 pct LPG, and 26.7 pct gasoline. THE coal hydrogenation process is well known as a means for the production of liquid fuels from coal. In this paper the possibilities of the coal hydrogenation process as a source of chemical raw materials will be discussed. The recovery of chemicals from coal hydrogenation products necessitates some modifications and additions to the equipment used for the production of fuels. The conversion of coal into liquid fuels such as gasoline requires two steps: liquid phase and vapor phase hydrogenation. Intermediate oils obtained in the liquid phase contain oxygen and nitrogen compounds, especially aromatic compounds, since the coal itself is composed mainly of cyclic compounds. These oils are the raw material for the separation of high molecular weight aro-matics and of tar acids and tar bases. The relative amounts of heavy and middle oil—boiling above and below, e.g., 325°C—can be influenced by temperature, residence time, and catalyst of the liquid phase hydrogenation. In the vapor phase hydrogenation a solid bed catalyst is used to convert the intermediate oil from the liquid phase operation into hydrocarbons of the gasoline boiling range. Essentially pure hydrocarbons with a low content of sulphur, oxygen, nitrogen, and olefines are obtained. Depending mainly on the catalyst, the high ring compound content of the liquid phase oil can be preserved and gasoline rich in aromatics obtained. However, with other catalysts, these ring compounds can be split into gasoline with a high content of isoparaffins.' Again, depending on the operating conditions, products of varying boiling range can be obtained. Selection of catalyst and operating conditions therefore makes possible the production of different grades of motor fuel and also variations in the relative amounts of benzene, toluene, and xylene when aromatic hydrocarbons are produced. The schematic flowsheet, Fig. 1, shows the main steps of the coal hydrogenation process, and the product streams used for the recovery of chemicals are indicated. In the liquid phase hydrogenation stage, coal is made into a paste with recycle heavy oil. The coal paste is mixed with hydrogen and preheated at a pressure of 300 to 700 atm in heat exchangers and preheaters. The coal paste and hydrogen enter high pressure converters where the liquefaction of coal, conversion of the high boiling liquid materials initially obtained, into intermediate oils is performed at a temperature around 470°C. The reaction products enter the hot catchpot in which lower boiling oil vapors are separated from the residue containing unreacted coal, ash, and heavy oil. This residue, usually called H.O.L.D. (heavy oil let down), is processed to separate solid and liquid material. By thermal treatment of the residue, e.g., a coking process, distillate oil is obtained, which can be recycled as part of the pasting oil. This stream is rich in polynuclear aromatics and can be, and has been, used for the recovery of such material. The oil vapors and hydrogen from the hot catchpot are cooled in heat exchangers and a cooler, and the condensed oil is separated from the recycle hy-

Jan 1, 1953

By E. E. Donath

Application of the coal hydrogenation process for the production of chemicals is described. It has been estimated that a plant to produce 31,090 bbl per day of chemicals and fuels would cost $326,- 000,000. These products would consist of 6.1 pct by weight tar acids, 50.8 pct aroma tics, 16.4 pct LPG, and 26.7 pct gasoline. THE coal hydrogenation process is well known as a means for the production of liquid fuels from coal. In this paper the possibilities of the coal hydrogenation process as a source of chemical raw materials will be discussed. The recovery of chemicals from coal hydrogenation products necessitates some modifications and additions to the equipment used for the production of fuels. The conversion of coal into liquid fuels such as gasoline requires two steps: liquid phase and vapor phase hydrogenation. Intermediate oils obtained in the liquid phase contain oxygen and nitrogen compounds, especially aromatic compounds, since the coal itself is composed mainly of cyclic compounds. These oils are the raw material for the separation of high molecular weight aro-matics and of tar acids and tar bases. The relative amounts of heavy and middle oil—boiling above and below, e.g., 325°C—can be influenced by temperature, residence time, and catalyst of the liquid phase hydrogenation. In the vapor phase hydrogenation a solid bed catalyst is used to convert the intermediate oil from the liquid phase operation into hydrocarbons of the gasoline boiling range. Essentially pure hydrocarbons with a low content of sulphur, oxygen, nitrogen, and olefines are obtained. Depending mainly on the catalyst, the high ring compound content of the liquid phase oil can be preserved and gasoline rich in aromatics obtained. However, with other catalysts, these ring compounds can be split into gasoline with a high content of isoparaffins.' Again, depending on the operating conditions, products of varying boiling range can be obtained. Selection of catalyst and operating conditions therefore makes possible the production of different grades of motor fuel and also variations in the relative amounts of benzene, toluene, and xylene when aromatic hydrocarbons are produced. The schematic flowsheet, Fig. 1, shows the main steps of the coal hydrogenation process, and the product streams used for the recovery of chemicals are indicated. In the liquid phase hydrogenation stage, coal is made into a paste with recycle heavy oil. The coal paste is mixed with hydrogen and preheated at a pressure of 300 to 700 atm in heat exchangers and preheaters. The coal paste and hydrogen enter high pressure converters where the liquefaction of coal, conversion of the high boiling liquid materials initially obtained, into intermediate oils is performed at a temperature around 470°C. The reaction products enter the hot catchpot in which lower boiling oil vapors are separated from the residue containing unreacted coal, ash, and heavy oil. This residue, usually called H.O.L.D. (heavy oil let down), is processed to separate solid and liquid material. By thermal treatment of the residue, e.g., a coking process, distillate oil is obtained, which can be recycled as part of the pasting oil. This stream is rich in polynuclear aromatics and can be, and has been, used for the recovery of such material. The oil vapors and hydrogen from the hot catchpot are cooled in heat exchangers and a cooler, and the condensed oil is separated from the recycle hy-

Jan 1, 1953

By H. Warlimont

In recent investigations of the microstructure and crystallographic features of martensite by electgon microscopy,', '9 thin films (about 50 to l000A in thickness) have been used as specimens. It was found by W. Pitschl ,2 that the lattice relationships between the supefiaturated fcc matrix and the martensite formed in thin films were different from those observed after transformation in bulk material. Corresponding differences are to be expected in nucleation and in reaction kinetics in the formation of martensite in thin films, but have not yet been reported. In the course of the preparation of thin films of iron-base alloys for investigation by electron microscopy it was observed that films of alloys, whose martensite formation temperature (Ms) was below room temperature, formed martensite during the electrochemical thinning process conducted at or above room temperature. The chemical compositions and approximate bulkM,-temperatures of the three materials on which this observation was made are listed in Table I. The alloys were vacuum-melted from high-purity metals. It can be seen that the effect occurred in alloys that contain only sub-stitutional components as well as in alloys that contain carbon as an interstitial solute. The appearance of martensite formed in this manner is illustrated in Fig. 1, which represents a polished film of alloy No. 1 at XI0 magnification. The martensite plates can be recognized by their surface relief effects. In addition, heavily strained (owrinkledo) austenite can be seen near the edges of the film where it is thinnest and transformation stresses or strains can be accommodated by plastic deformation most easily. The nonuniform distribution of transformed regions is probably due to the varying thickness of the film. These observations were common to all materials listed in Table I. Because effects of supercooling or straining can

Jan 1, 1962

By Zdzislaw Specht

THIS paper discusses elementary laws pertaining to seismic reflection methods and presents a general and simple criterion for determining the direction of dipping of a reflection horizon from observational data of distance and time. The important development of practical geophysics and the successful results obtained in America, Germany, and other countries, caused introduction of geophysical methods of prospecting in Poland. Of all methods, seismic prospecting has proved most successful in the area of the Carpathian foreland, where new oil reserves are to be expected. Because of the great depth of the sedimentary column in that area, rang-ing sometimes over 2000 meters, it was obvious that only the seismic reflection method could give reliable information regarding the structural features of these strata at such a depth. During the experimental surveys it appeared that the complicated geological conditions of that area do not always allow prediction of the depth and the angle of dipping of reflecting horizons, and for that reason information about the inclination of a given horizon has great importance in the determination of subsurface structure. Therefore the problem arises of obtaining from seismic reflection data reliable information as to the direction of dipping of reflecting strata, without any previous knowledge of the average velocity and depth of geologic horizons under consideration.

Jan 1, 1940

By Edward Hart

The specimen of glazy iron used for analysis was highly characteristic in appearance. It was made at one of the furnaces of the Glendon Iron Works, working a light burden of ore with a highly silicious cinder. There are very few records of analyses of this variety of iron. I have been able to find but the two given by Bell, and I add these to my analysis for comparison. Bell. 1. 2. Iron,......89.47* 88.18 90.70 Carbon, combined,.... 0.00 0.79 0.71 Graphite,......2.88 2.59 2 68 Silicon,......5.47 5 13 5.13 Manganese,.....1.54 0.77 0.56 Sulphur,......0.04 0.17 0.23 Phosphorus,.....0.60 1.12 1.12 Titanium,.....- 0.26 0.18 Calcium,......0.00 0.22 0.20 Magnesium,.....0.00 0.06 0.03 100.00 9929 101.54

By L. A. Panek, F. D. Wang, M. C. Sun

A method of stability analysis of underground opening by the limiting equilibrium approach is presented. The Coulomb shear failure criterion is employed to find possible fracture surfaces and to evaluate the shear strengths along them. The normal stresses required for the calculation of shear strength, and the shear stresses for the calculation of total shear force along the fracture surfaces, are obtained by the finite-element method of plane-strain analysis. The factor of safety against Coulomb shear failure is calculated as the ratio of the shear strength to the shear force along a fracture surface, similar to the practice in slope stability analysis. Four shapes of openings are analyzed, with three ratios of horizontal lithostatic stress and four coefficients of internal friction. The critical fracture surfaces found by this method of analysis are consistent with the arching of roofs and slabbing of ribs commonly observed.

Jan 1, 1972

By R. M. Goldhoff, J. W. Spretnak, R. Speiser

IT has been observed by Grange and Garvey' that the homogenization of boron-treated steels could lead to complete elimination of the hardenability effect caused by boron. The experimental conditions leading to this conclusion involved the heating of steel specimens with and without boron at 2350°F for 24 hr. To prevent oxidation and decarburization, the specimens were encapsulated under vacuum in silica tubes. This observation of permanent deterioration of boron effect was noted in both a 0.63 pct plain carbon steel and a 0.25 pct C low alloy steel, both of which had been treated with ferroboron and homogenized at 2350°F for 24 hr. In another series of boron-treated steels which had been homogenized for shorter times, the effect was to reduce, but not completely eliminate, the boron effect. Under these circumstances, it was presumed that longer heating times would cause permanent deterioration of the boron effect in this series of steels also. Since the characteristic grain boundary constituent could not be formed in any of these steels after homogenization, it was concluded that possibly boron was being converted into an ineffective form through chemical combination resulting from prolonged heating at high temperatures. Assuming the validity of this observation, a series of experiments to establish the relationship between temperature of homogenization and kinetics of deterioration was contemplated. Initially an attempt was made to reproduce the results just described. For this purpose commercial lots of AISI 8640 and AISI 86B40 steels were used. The analyses of these steels appear in Table I. Hardenability evaluation was on the basis of the hardness of normalized sections. Grange and Garvey' pointed out that these steels are air hardening to some extent and such a technique can be used. For these tests the specimens were held for 20 min in a salt bath at 1600°F and then cooled in still air. Several evacuated silica capsules containing pairs of the test steels were made up and heated for 24 hr at 2350°F. A number of failures occurred among these capsules, but evaluation of the specimens successfully treated indicated a complete deterioration of the boron effect. However, examination of the surface of the treated specimens showed that some reaction, probably with the silica tube, had taken place, since none of the specimens was bright. One further test was made at 2350°F for 12 hr, using a pair of specimens wrapped in tantalum foil prior to encapsulation. These specimens appeared bright after treatment, and their hardenability evaluation showed that no deterioration had taken place. Since these steels are extremely sensitive to de-boronization at high temperatures and particularly at long times of holding, the experimental observations raised some doubt as to the real occurrence of the permanent deterioration effect. To resolve this question an experiment was undertaken in which the high temperature portion of the treatment was completed in vacuum. For this purpose a molybdenum wound resistance furnace capable of maintaining good vacuum was used. All the pertinent accessories to this furnace were machined from molybdenum stock in order to withstand the high temperature involved in the experiment. With this equipment a temperature of 2400 °F was obtained and held under a pressure of 5x10-" mm Hg for 24 hr. A specimen of each of the two test steels was suspended in the furnace and both were heated simultaneously. Temperature was measured with a standardized Pt—Pt-13 pct Rh thermocouple, whose millivoltage at 2400°F was extrapolated from the standardization equation, good up to 2200°F. The complete evaluation of these specimens is shown in Table 11. The hardness values attained by each steel after normalizing, both before and after the prolonged high temperature treatment, remain the same. A microscopic study of the decomposition products present in the microstructures of specimens at various points in the heat treatment correlated with the hardness values observed. It seems that no permanent loss of boron hardenability effect occurs, at least in these Grainal-treated steels, when deboronization is circumvented. Acknowledgments This work was conducted under the sponsorship of the Aeronautical Research Laboratory, Wright Air Development Center, under Contract No. AF 18(600)-94, and their support and permission to publish is gratefully acknowledged. A note of thanks is also extended to the Republic Steel Corp. for supplying the materials used in the work. References ' R. A. Grange and T. M. Garvey: Factors Affecting the Harden-ability of Boron Treated Steels. Tro.ns. ASM, 1946, vol. 37, p. 136.

Jan 1, 1957

By Joseph Moses Botbol

In passing the judgments essential to his trade, the geologist may be assisted by two new methods. These will enable him to (1) arrange the minerals of a selected group of base metal mining districts in the order of their typicalities and (2) compute the degree of mineralogic association between the model so derived and a district external to the model. It must be remembered, of course, that the methods discussed in this article are no more powerful than the structure of the model used and the data available. The geologist who employs these techniques must still exercise his expertise

Jan 1, 1970

By Thomas H. Sheldon

IN the zeal for opening up new ore-bodies, or for. extracting the ore from attractive bodies gal ready opened up, we very often lose sight of the fact, that, after all, the operation of a mine is a business proposition, pure and simple; and for the best working-results should be treated upon a strict business basis. Of course, in every mine of consequence a record is kept of expenditures and receipts, and such glittering generalities as "gross receipts," " net receipts," " mining expenses," .and " per cent. profit," can be told to the cent; but does this . record show economy of management, either as compared with the same record of other months, or as compared to the record of other mines of the' same class ? Moreover, if such a record shows that the cost of mining is high, does it in any way en¬able the manager to put his finger on the leakage? Does it necessarily follow that a mine which makes a profit of, say, 40 per cent on 25-dollar ore, is doing less economical mining than one which saves 60 per cent on 80-dollar ore ? Of course the figures in the latter case look the more attractive, but when it comes to a point of saving everything which can be saved, and of cutting down expenses to the lowest possible cost of operation, the former mine is doubtless on the most firm and economical financial basis. But as the relative merits of the system of mining in the two cases, nothing could be decided without a basis of detailed comparison ; if one system is more economical than the other, why is it so, and wherein does the advantage lie ? This can be shown only by keeping accurately the cost of each mining operation. And no matter how dissimilar two mines may be in character and operation, yet there are a few general heads common to all mining operations. In the first place, it is necessary to break the ore from the solid ground ;

Nov 1, 1905

By L. Adler

The proper transfer of roof loads from props and bolts to ribs and pillars can result in appreciable savings. The author shows how to plan such load reduction in underground mines. For openings in bedded rocks, analyzed by simple beam theory, it has been shown that roof loads can be shifted from one support to another.&apos; This transfer is effected by controlling the relative deflection of adjacent supports. If this action is used to reduce loads in the props or bolts and increase them in the ribs or pillars, it could be productive of economies averaging 30 pct. Such savings can be realized by: 1) using smaller artifical supports such as props and bolts, and 2) increasing loads in the rib to induce eventual failure for longwall caving. Since the rib is relatively load-insensitive, it provides a natural load sink. However, caution must be used in applying such permissive deflections, for while the supports are being handled properly, the stresses in the roof itself are very sensitive to such manipulations.1 An underground opening overlain by a relatively thin immediate roof, and a thick main roof, is usually laid out as shown in Fig. 1. The total open span (T) between pillars or ribs is calculated on the basis of the main roof, and the secondary spans (L) on the basis of the immediate roof, so that a complete accounting of the existing loads is made. For this purpose the immediate roof is treated as a uniformly loaded continous beam, which is an indeterminate structure as shown in Fig. 2(A). The term systematic supports places the following two restrictions on the continuous beam: 1) Spans between all supports are equal: uniformity of spans. 2) Loads on all supports (excepting ribs) are equal: uniformity of supports. The usual mode of analysis of such a structure by beam theory is to divide it up into a series of simply supported beams with end moments as shown in Fig. 2(B)2 The problem now is to transfer a maximum roof load to the rib from the span immediately adjacent to it by permitting its support to deflect, and yet not fail the roof. Once this has been done, the load can be equalized on all other supports by permitting further degrees of relative deflection. It is necessary, therefore, to determine an expression for the constant load on each support, and the required deflections, in terms of properties of the roof. From Fig. 1: T/L0 = No and T/L = N [1] Where: T = allowable span of main roof, ft. (Calculated from S = Mc/l) S is allowable modulus of rupture, psi. M is maximum moment, lb-ft (in this case wT2/12). c is distance from neutral axis to extreme fiber, ft. I is moment of inertia with respect to the neutral axis, ft4. L,, = allowable span of immediate roof, ft (calculated as was T). L = actual span of immediate roof, ft. N0 and N are pure numbers where No < N and N is the integer just higher than Nu. For example: If T =100 ft and L0 - 15 ft 100 N0 =100/15= 6.667 . . .N= 7 L = T/N = 100/7 = 14.28 ft Consequently N = No + n where n is a positive fraction i.e.: 0< n < 1 Since from Eq. 1, T = NL = N0L0 Lo /L = N/No = N/N-n

Jan 1, 1961

By H. P. Willard, R. W. Gates

Basic Inc., has been mining magnesite and brucite for 27 years at Gabbs, Nev. This is a medium sized open-pit mine. The magnesite deposits are in north-western Nye County about 30 miles east of Luning, Nev., on the west slope of the Paradise Range, at an elevation between 5000 and 6000 ft. The magnesite and associated brucite is used to make magnesia refractories employed principally in the steel industry and magnesium oxide for various chemical uses.

Jan 10, 1963

By Ingvar Janelid

DURING the last ten years, in the effort to save manpower and costs, methods of drilling and blasting in Sweden have changed and developed in a revolutionary manner. These developments have been accompanied by extensive alterations in methods of tunneling and in mining technique. Since the end of the 1940&apos;s there has been a 100 pct change-over in Sweden to the use of tungsten carbide tipped drill steel. Both investigations and practical results have shown clearly that a tungsten carbide chisel bit fastened directly on the drill steel is the most economical for Swedish conditions. This applies to drill steel up to the longest size it is practical to handle. In terms of drilling footage, durability of the drill steel and the tungsten carbide bit is generally the same. In this respect, therefore, a detachable bit offers no advantage over integral steel. Transportation of integral steel to the grinding plant, instead of detachable bits only, involves heavier work, but this is counteracted by placing grinding plants closer together or by grinding at the working place itself. The drilling rate with integral steel is higher, the trouble of changing bits is avoided, and the difficulties which always arise with joints are eliminated. At present steel dimensions are usually 7/8 in. with a shank 41/4 in. long, whereas the international standard shank was formerly 31/4 in. The longer shank renders possible a better guide control of the steel and consequently longer life for the rotation chuck and the water-flushing pipe. After extensive experiments and investigations, agreement has been reached in Sweden on certain series of standards, shown in Table I. Other steel dimensions are, of course, available for special purposes. The standard drill-steel set begins with a 34-mm diam and has a length interval of 2 2/3 ft. On the other hand, for ordinary drifting it has been proved more economical to use greater length intervals on the steel, so only one steel change is required for a 61/2-ft round. Moreover, owing to the relatively close hole spacing necessary in a drift round, it has been possible to reduce the diameters of the tungsten carbide bits. This results in faster and cheaper drilling. As shown by the table, the lengths of drill steels for normal drifting are 33/4 ft and 71/2 ft, with 29 and 28-mm diam bits. During the last few years operators have begun to use 3/4-in. drill steel in place of the 7/8-in. drill steel chiefly employed earlier. This is a valuable development in drilling technique, particularly in view of

Jan 1, 1955