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By John C. Cosgrove

IT was just five minutes past midnight, on Wednesday, Jan. 5, 1938, that Mrs. Cosgrove and I sailed from New York City. Our trip was to completely circle the globe, to cover over 40,000 miles and stop at thirty ports-from New York to Trinidad, to South America, via South Africa to the Orient, to India, to the Malay States, to Indo-China, the Philippines, Australia, New Zealand, Fiji, and across the Pacific home via Honolulu. It is not my intention to make this a complete, technical study of coal or coal mining in the various countries that .we visited but more to outline the general in- formation that the mining engineer and coal operator would absorb in his travels. Coal is the universal fuel and is in use wherever there is commerce and industry. Even on our ship, which, like practically all ocean liners was an oil-burner, a considerable quantity of coal was used for cooking.

Jan 1, 1939

By F. W. Osterwald, C. R. Dunrud

Coal mine bumps are a serious hazard to life and property in the mines of east-central Utah. Research into geologic factors associated with these bumps indicates that the bumps are spatially and genetically related to structural and stratigraphic features. Some bumps are directly related to stress accumulation along faults, either from natural causes or as a result of mining. Frictional properties of bedding planes between coal and roof rock, and between some rock units within the roof, directly affect the deformation of coal ribs, and hence the incidence of bumps. These frictional properties are related to the lithology of the roof rocks as well as to the sedimentary structures within the rocks. Commonly a sequence consisting of 1/3 ft of carbonaceous silt-stone, about 2/3 ft of rider coal, and several ft of massive siltstone or sandstone overlies the main seam. The sequence causes difficult roof conditions in the mines, difficult mining conditions and, directly or indirectly, many bumps. The U. S. Geological Survey has studied the relationships between geological features and coal mine bumps at Sunnyside, Utah, for about six years. The work, which was originally undertaken at the request of and in cooperation with the U.S. Bureau of

Jan 1, 1965

By W. N. Poundstone, P. C. Thakur

Introduction Horizontal drilling in coal mines is a relatively new technology. The earliest recorded drilling in the United States was done in 1958 at the Humphrey mine of Consolidation Coal Co. for degasification of coal seams. Spindler and Poundstone experimented with vertical and horizontal holes for several years. They concluded in 1960 that horizontal drilling in advance of underground mining appeared to offer the most promising prospect (for degasification) but effective and extensive application would be dependent upon the ability to drill long holes, possibly 300 to 600 m, with reasonably precise directional control and within practical cost limits (Spindler and Poundstone, 1960). Mining Research Division of Conoco Inc., the parent company of Consolidation Coal Co., began a research program in the early 1970s to achieve the above objective. The technology needed to drill nearly 300 m in advance of working faces was developed by 1975 and experiments on advance degasification with such deep holes began in 1976. Preliminary results of this research have already been published (Thakur and Davis, 1977). To date nearly 4.5 km of horizontal holes have been drilled for advance degasification and earlier results were reconfirmed. In summary, these are: • The greatest impact of these boreholes was felt in the face area where methane concentrations were reduced to nearly 0.3% in course of two to three months from original values of nearly 0.95%. • The methane concentration in the section return reduced to 50% of its original value immediately after the boreholes were completed, indicating a capture ratio of 50%. • The total methane emission in the section (rib and face emission plus the borehole production) did not increase but rather gradually declined with time. • Initial production from 300 m deep boreholes in the Pittsburgh seam varied from 3 m3/min to 6 m3/min but then slowly declined as workings advanced inby of the drill site (well head) exposing a larger surface area parallel to the borehole. Encouraged by these results, it was decided to design a horizontal drilling system that would be mobile and compatible with other face equipment. A mobile horizontal drill can be divided into three subsystems: the drill rig, the drill bit guidance system, and borehole surveying instruments. The drill rig provides the thrust and torque necessary to drill 75- to 100-mm diam holes up to 600 m deep and contains the mud circulation and gas cuttings separation systems. The drill bit guidance system guides the bit up, down, left, or right as desired. Borehole surveying instruments measure the pitch, roll, and azimuth of the borehole assembly. Additionally, it also indicates the thickness of coal between the borehole and the roof or floor of the coal seam. Thus, it becomes a powerful tool for locating the presence of faults, clay veins, sand channels, and the thickness of coal seam in advance of mining. In recent years, many other potential uses of horizontal boreholes have come to light, such as in situ gasification, longwall blasting, improved auger mining, and oil and gas production from shallow deposits. The purpose of this paper is to describe the hardware and procedure for drilling deep horizontal holes. The Drilling Rig [Figures 1 and 2] show the two components of the mobile drilling rig: the drill unit and the auxiliary unit. The equipment (except for the chassis) was designed by Conoco Inc. and fabricated by J. H. Fletcher and Co. of Huntington, WV. The drill unit. It is mounted on a four-wheel drive chassis driven by two Staffa hydraulic motors with chains. The tires are 369 X 457 mm in size and provide a ground clearance of 305 mm. The prime mover is a 30-kw explosion-proof electric motor which is used only for tramming. Once the unit is Crammed to the drill site, electric power is disconnected and hydraulic power from the auxiliary unit is turned on. Four floor jacks are used to level the machine and raise the drill head to the desired level. Two 5-t telescopic hydraulic props, one on each side, anchor the drill unit to the roof. The drill unit houses the feed carriage, the drilling console, 300 m of 3-m-long NQ, drill rods, and the electric cable reel for instruments. The feed carriage is mounted more or less centrally, has a feed of 3.3 m, and can swing laterally by ± 17°. It can also sump forward by 1.2 m. The drill head has a "through" chuck such that drill pipes can be fed from the side or back end. General specifications of the feed carriage are: [ ] The auxiliary unit. The chassis for the auxiliary unit is identical to the drill unit but the prime movers are two 30-kW explosion proof electric motors. It is equipped with a methane detector- activated switch so that power will be cut off at a preset methane concentration in the air. No anchoring props are needed for this unit. The auxiliary unit houses the hydraulic power pack, the water (mud) circulating pump, control boxes for electric motors, a trailing cable spool, and a steel tank which serves for water storage and closed-loop separation of drill cuttings and gas.

Jan 1, 1981

By L. O. Brockway, W. C. Bigelow, J. A. Amy

Specimens of Inconel-X alloy solution-treated at 2050°F and aged for periods of 1, 10, 100, and 1000 hr at 1200°, 1400°, and 1600°F have been examined by electron microscopy and by electron and X-ray diffraction. Four minor phases were identified: titanium nitride, niobium nitride, an M23C6 complex carbide, and the intermetallic ? phase. The two nitrides were found in all of the aged specimens and in the solution-treated alloy, the y' phase formed at all aging temperatures, while the M23C6 carbide formed only on aging at 1200° and 1400°F. Variations in the hardness, rupture strength, creep strength, and yield strength of the alloy have been shown to be related to variation in the size and distribution of the particles of the y' phase within the matrix grains of the alloy, with maximum values of these properties occurring when there is a maximum amount of the phase present and a fine state of dispersion of the particles prevails. THE nickel-chromium alloys containing titanium and aluminum are widely used in applications requiring high strength at elevated temperatures. Studies on relatively simple ternary and quaternary alloys'-4 have indicated that the favorable high-temperature properties of these alloys are related to the precipitation of an intermetallic ?'-phase having a structure based On that of the Ni3A1 phase of the nickel-aluminum system. The present investigation was undertaken to identify the minor phases and to study the microstructural changes occurring in one of the more complex commercial alloys as a result of aging in the service temperature range, with particular emphasis on investigating the role of the ?' -phase. The microstructural changes were studied by electron microscopy and the minor phases were identified by X-ray and electron diffraction. MATERIALS AND METHODS Inconel-X alloy was selected for this investigation because the previous studies of Frey, Freeman, and white6 provided extensive information on the variations in metallurgical properties of this alloy resulting from high-temperature aging treatments. Specimens aged 1, 10, 100, and 1000 hr at 1200°, 1400°, and 1600°F were made available for the present use. These were taken from 1-in. round-bar stock of a commercial heat having the following weight composition: 0.04 C, 0.56 Mn, 6.59 Fe, 0.007 S, 0.38 Si, 0.03 Cu, 73.22 Ni, 14.97 Cr, 0.78 Al, 2.38 Ti, and 1-02 Nb. They were solutjon-treated at 2050 for 4 hr and then water-quenched prior to aging. Powder X-ray diffraction patterns were obtained from insoluble minor phases which were separated from the alloys by digestion in the reagent of Mahla and Nielsen.6 Back-reflection patterns were obtained from electrolytically polished surfaces of the bulk search to detect the ?'-phase' Cu-K a radiation was used in both cases. In the electron-diffraction studies, the general method developed by Heidenreich, Sturkey, and

Jan 1, 1959

By E. I. Salkovitz, F. W. von Batchelder

AS Slawson1 and others have already shown, from a geometrical view twinning in the diamond structure may be considered as a reflection across a (111) plane or as a reflection across a (112) plane. An analysis by Ellis and Treuting2 showed that the twinning plane must be the (111) plane, but that a twin might have lateral boundaries along (112) planes. Earlier Slawson1 and Ellis3 had shown that only (111) twinning was present in diamond and germanium respectively. Since an ingot of silicon, 1 in. in diam and 1 1/2 in. long, prepared from Dupont silicon to which 0.2 pct Be had been added was available, it was examined for twinning. The ingot had been cooled slowly from the bottom. A section 1/16 in. thick was ground off from the bottom and then the ingot was sliced vertically through the center. Spectrographic analysis indicated only faint traces of iron and silver, in addition to the beryllium. The sample was first etched with a hot solution of 10 pct NaOH in water.and photographed. Back-reflection Laue patterns were made of the vertical and horizontal sections in the regions designated in Fig. 1. Orientations were obtained for each twin using the Greninger technique. Several regions of the same orientation were found; thus regions A and K were identical, similarly N and P, and C, G, J, and L. Knowing the angles between the traces of the twinning plane on the surface and the specimen axis, the indices of the twinning plane were deduced. In all cases the twinning plane was found to be (111); lateral boundaries, however, were found which were (112) such as the one indicated in Fig. 1. Therefore, as in the case of diamond1 and germanium' only (111) twinning occurs in silicon. We wish to thank Mr. Hino for the use of the ingot as well as for its metallographic preparation. References 1C. B. Slawson: American Mineralogist (1950) 35, p. 193. 2 W. C. Ellis and R. C. Treuting: Trans. AIME (1951) 191, p. 53; Journal of Metals (January 1951). 3W. C. Ellis: Trans. AIME (1950) 188, p. 886; Journal of Metals (June 1950).

Jan 1, 1953

By J. R. Eckel

It has long been known that mud properties affect drilling rate and that drilling with water is as much as six to .seven times faster than with mud. However, it is not known why drilling rates with mud are lower. Recent studies show what fluid properties govern microbit drilling rate and also provide a quantitative correlation of these fluid properties and hydraulics with microbit drilling rate. All the microbit tests reported in this paper were made with a low permeability (about 10 md) limestone. They were conducted at constant bit weight and rotary speed with varying fluid properties, flow rate Q and nozzle dinnteter d. The tests showed that (1) drilling rate in a given system with constant circulating rate and nozzle velocity is a function of the kinematic viscosity [viscosity /density, (cp)/ p (density)] of the drilling fluid measured at near bit nozzle shear rates; (2) the combined effect of fluid properties and hydraulics on microbit drilling rate is defined by a Reynolds number function (kQp/dp); (3) for the same kinematic viscosity, drilling rate is independent of solids content; and (4) for the same kinematic viscosity, drilling rate is independent of fluid loss. INTRODUCTION The fact that mudding up reduces rate of penetration is well known, but the reduction varies from well to well. The rate with mud may be 80 percent of the water rate or it may be as low as 15 percent. Since this reduction of up to seven-fold takes place under conditions of constant bit weight, rotary speed and mud circulating rate in a uniform formation, it can only be a result of the change in fluid properties. Previous work1-" both in field tests and in laboratory microbit tests) showed that drilling rate with mud is lower than with water (Figs. 1 and 2). Microbit tests also confirm, at least qualitatively, the field observation that drilling rate slows with increased pressure differential from borehole to formation (Fig. 3). The qualitative effect of drilling rate of bit weight, rotary speed and rate of mud circulation are also the same for the field and microbit rig. For these reasons, the industry generally accepts microbit results as being qualitatively meaningful in terms of field application.

By C. H. Jr. Herty

IT Is well known that manganese will desulfurize molten iron through the formation of manganese sulfide, which, being only slightly soluble in the metal, rises to and enters the slag where it remains as such (mixer slag) or is converted partly or wholly to calcium sulfide (basic open¬hearth slag). The reaction by which manganese sulfide is formed is Mn + FeS = MnS + Fe. This reaction is generally considered to be reversible, although McCance1 states- that such is not the case. The experimental work was carried out to confirm the results of Rohl2 as to the reversibility of the reaction. Six trials are herein considered. In each 200 grams of electrolytic iron free from manganese and sulfur were melted in a pure zirconia crucible in a carbon resistance electric furnace. To the melted iron in runs 2, 3 and 4 metallic manganese and manganese sulfide were added to give 1, 2 and 4 per cent. manganese respectively, with 0.50 per cent. of sulfur in each. In runs 5, 6 and 7, metallic manganese andiron sulfide to give the same proportions of manganese and sulphur as in runs 2, 3 and 4, were added. The melts were held molten in the furnace for an hour and then cast into small ingots.

Jan 2, 1925

By Elwin L. Fisk

The Cordero mine is located in the Nevada portion of the Opalite Mining district that straddles the Oregon-Nevada state boundary. The mercury deposits of the district occur along the margins of the broad east-west graben valley of McDermitt Creek. The only other major producers, the Opalite and Bretz mines, on the northern margin of the graben have thus far exploited only shallow deposits within 100 feet of the surface. Through December 1965, Cordero has produced 94,680 flasks of mercury, the most of which came from the large underground cinnabar deposit. Nearly all of the mercury ore deposits in this district have been found in Miocene rhyolitic intrusives, flows, tuffs, and lake beds. These rhyolites have been argillically altered at depth and extensively silicified at or near the surface. These silicified areas or "opalite" masses were the original cinnabar-bearing outcrops.

Jan 1, 1968

By William A. Longacre

In a previous paper the author discussed the problem described by the title of this paper, outlining and reviewing the Gish-Rooney method, with comparison and brief analysis of the interpretation techniques of Roman2 and Tagg.3 Some aspects of the simple type of two-layer system were reviewed from various sources.4, 5, 6 The Two-layer Problem For the configuration shown in Fig. I, and for any electrode separation A, the SurfAce of earth p, DepthH i P2 Infinite depth Fig. i.—Two-Layer system. apparent resistivity is a function of the relative depth the overburden resistivity pl, and a factor K defined by the equation P2 + P1 where p, is the resistivity of the lower bed. Thus, by the known theory of electrical images, p = pi [1+ */(*-7)] where The Multilayer Problem As the number of isotropic, homogeneous, horizontal layers increases, the difficulties of interpretation multiply. The theory becomes much more involved, and in view of the state of our knowledge, the general problem is virtually impossible of exact solution. However, certain types of multilayer problems can be worked out in terms of two-layer theory. Humme1,7 Pirson,8 Tagg,3 Watson,9 and others have pointed out the possibility of building up approximations of multilayer resistivity curves with suitable combinations of two-layer curves. Many three-layer curves can be constructed in this way to a good degree of accuracy, and special types of four-layer curves can be approximated. It is very likely that as the number of layers increases, the method of combination will not give good approximations except in very special cases. In the analysis of actual field data, the author has found that many obviously multilayer curves can be solved partially with a good degree of accuracy, using the theory that thin adjacent beds may be considered as one layer or zone, and applying two-layer theory to the extremities of the field curve. Modified Tagg Method of Interpretation Tagg's second method, though superior to his first, necessitates a large number

Jan 1, 1946

By William A. Longacre

In a previous paper the author discussed the problem described by the title of this paper, outlining and reviewing the Gish-Rooney method, with comparison and brief analysis of the interpretation techniques of Roman2 and Tagg.3 Some aspects of the simple type of two-layer system were reviewed from various sources.4, 5, 6 The Two-layer Problem For the configuration shown in Fig. I, and for any electrode separation A, the SurfAce of earth p, DepthH i P2 Infinite depth Fig. i.—Two-Layer system. apparent resistivity is a function of the relative depth the overburden resistivity pl, and a factor K defined by the equation P2 + P1 where p, is the resistivity of the lower bed. Thus, by the known theory of electrical images, p = pi [1+ */(*-7)] where The Multilayer Problem As the number of isotropic, homogeneous, horizontal layers increases, the difficulties of interpretation multiply. The theory becomes much more involved, and in view of the state of our knowledge, the general problem is virtually impossible of exact solution. However, certain types of multilayer problems can be worked out in terms of two-layer theory. Humme1,7 Pirson,8 Tagg,3 Watson,9 and others have pointed out the possibility of building up approximations of multilayer resistivity curves with suitable combinations of two-layer curves. Many three-layer curves can be constructed in this way to a good degree of accuracy, and special types of four-layer curves can be approximated. It is very likely that as the number of layers increases, the method of combination will not give good approximations except in very special cases. In the analysis of actual field data, the author has found that many obviously multilayer curves can be solved partially with a good degree of accuracy, using the theory that thin adjacent beds may be considered as one layer or zone, and applying two-layer theory to the extremities of the field curve. Modified Tagg Method of Interpretation Tagg's second method, though superior to his first, necessitates a large number

Jan 1, 1946

By George C. Gibbons

ALMOST everyone in any of the five counties embracing the great East Texas field depends heavily upon oil for his living whether or not he actually owns a well or piece of royalty himself. Oil is a natural resource that, when utilized, spreads its benefits over a vast area and into many channels. Depending then almost entirely upon agriculture: what is now known as the East Texas oil field a decade ago had a scattered farming population. For many of its people, tomatoes and other vegetable crops coupled with the old southern stand-by of cotton, furnished an adequate living and a small cash surplus. Some farm families on the red hills, however, were not far re- moved from the sowbelly and black- eyed peas, or cornpone and potlikker, standard of living. Kilgore of that day, was a village of less than 850 per- sons. Longview, with just over 5000, apparently had reached the limit of its growth. Tyler was an overgrown farming town of around 17,000 persons, de- pending largely upon the agricultural income of Smith County and surrounding territory for which it was a trading center.

Jan 1, 1941

By H. C. Fowler, G. B. Shea

MODERN industry's incessant demands for increased operating efficiency and lower costs require that hazards attending all occupations be reduced to a minimum. Reduction of the inevitable losses that result from accidents and injury to the worker is one of the important economic problems confronting industry today and is the basis for the widespread safety movement. Dr. Charles E. Munroe, chief explosive chemist of the U. S. Bureau of Mines, is father of the statement, "The existence of a nuisance is evidence of an economic waste." This pointed expression of the internationally recognized scholar and explosives engineer has a direct relation to the prevention of accidents. Every injury and death in industry is a waste in human endeavor and a lessening of productive gain. Dr. Munroe has aptly paraphrased his previous summation with reference to wastes as follows: "A lost-time accident is evidence of an economic waste." That the economic waste entailed by occupational accidents is chargeable to industry is well substantiated by the fact that all but four states (Mississippi, South Carolina, Florida and Arkansas) now have workingmen's liability and compensation laws.

Jan 1, 1933

By A. W. Bethune, L. M. Pidgeon

The equilibrium vapor pressure of zinc has been determined over the systems: ZnS(s) + Fe(s) = FeS(s) + Zn (vapor) and ZnS(s) + 2Cu(s) = Cu2S(s) + Zn (vapor) by reacting the components in an evacuated tube containing a thin copper fiber, and allowing equilibrium to be established with respect to the brass formed and the zinc in the vapor phase. The composition of the brass formed was determined and equilibrium vapor pressure values obtained from existing data. Between 850' and 1000°C, values for the first reaction ranged from 8 to 58 mm Hg; for the second reaction from 1.8 to 22 mm Hg. THE direct reduction of metallic sulphides may be indicated by an equation of the form MS + X = XS + M [1] where X is some suitable reducing agent. In the case where M is a metal of relatively high volatility, the use of vacuum will displace the equilibrium to the right at comparatively low temperatures in the face of the usually unfavorable thermochemistry. For this technique to be feasible, MS, X, and XS, must of course be nonvolatile compared to M. In this regard, zinc becomes an obvious metal both from the point of view of its volatility and the fact that most commercial ore deposits of this metal occur as the sulphide, and its direct reduction would avoid the expensive roasting process necessary prior to present day reduction methods. Two reducing agents are of interest, namely copper and iron, with the former having the added attraction of being relatively easily regenerated by oxidation in the copper converter. The possibility of these reactions has been considered before. Imbert1 patented a process whereby zinc sulphide was reduced by either copper or iron. The process as described took place at atmospheric pressure and, at the high temperatures required, the system was molten, resulting in the formation of a matte with the consequent reduction in the activities of the reactants. Peterson2 did further experimental work on these reactions and, although promising results were obtained on a small scale, attempts to increase the size of the furnace led to excessive "blue powder" formation. The early workers failed to realize the large reduction in temperature afforded by the use of vacuum and its desirable results. More recently, the direct reduction of zinc sulphide by iron has been the subject of a thermody-namic investigation by Kelley,3 and Gross and War-rington' have studied the kinetics of the reaction in the temperature range 900" to 1000°C, their experiments being conducted using vacuum and indicating that the reaction would proceed practically to completion in 1 to 3 hr, depending on the temperature. A qualitative examination of the system ZnS-Cu carried out at this University indicated that 98 pct of the theoretical zinc in the charge could be recovered in 1 hr at 1000°C when operating under reduced pressure. The purpose of this research program was to study the equilibrium of the two reactions. During the investigation, reference to the experimental work of Schenk5 was found but, due to lack of detail in the report, it was felt that it would be desirable to continue the equilibrium measurements. Selection of Method The measurement of vapor pressure of metals presents obvious problems resulting from the high temperatures which must be employed. In the present case, the production of the "reaction pressure" introduces a further difficulty since a chemical reaction must proceed before any metal vapor pressure is produced, resulting in depletion of the re-

Jan 1, 1954

By R. Wang, N. J. Grant, B. C. Giessen

By crystallographic and X-ray methods, the existence and isonzorphism of Ti3Rh5 and Hf3Rhs were confirmed. Both phases are of the orthorhombic Ge3Rh5 type; lattice parameters and refined positional parameters are given. The structure is related both to the filled-up NiAs-B8 and Cu-AI types. An analogous phase with zirconium does not exist; the effect of ternary substitutions for titanium ad hafnium suggests a size factor limit to be active. A recent survey of phase diagrams of the T4 metals titanium, zirconium, and hafnium with the T, noble metals rhodium and iridium indicated the existence of the A3B5 phases Ti3Rhs, ZrsRhs, and HfsRhs. Ti3Rhs and Hf3Rh5 were found to be isostructural, based on the line-rich powder patterns which had not been analyzed. Zr3Rh5 was considered to have a substructure of the NbRu type (orthorhombically distorted B2-CsCl type).&apos; Because, in combinations with other transition metals, hafnium and zirconium are generally more likely to form isostructural phases than hafnium and titanium (with the significant exception of the Ti2Ni-"E93" type phases based on T4 metals2), the reversal of this relation for the A3B5 phases was of interest. As shown in the following, the nonexistence of Zr3Rhs has been established, the structures of Ti3Rh5 and Hf3Rh5 have been worked out, and crystal chemical relationships and stability criteria are reported. EXPERIMENTAL METHODS AND RESULTS Alloy Preparation and Phase Diagram Work. Alloys were prepared from high-purity (99.99+ pct) elements by arc-meltin3,4.Heat-treated alloys were annealed in a vacuum of 3 x X torr for 24 hr at 1300DC. Metal-lographic samples were etched electrolytically in concentrated HCl with 5 v ac for 5 min.3 X-ray diffraction powder patterns were taken on a GE XRD-5 dif-fractometer with Cum radiation at low scanning rates (0.2 deg per min for 28). It was confirmed that Ti3Rhs and Hf3Rh5 have similar diffraction patterns, and that an alloy with the composition Zr3Rhs has a different pattern. Six Zr-Fh alloys with 59 to 69 at. pct Rh were therefore prepared and investigated in the as-cast state by X-ray diffraction and metallography. Alloys at 59 and 61 at. pct Rh were found to be a single phase, with the distorted B2-CsC1 type structure typical for the off-stoichio- metric region of the phase (Zr,-,Rh,)Rh. This phase forms a eutectic with ZrRhs at about 66 at. pct Rh: accordingly, alloys between 61 and 69 pct at. pct Rh consisted of two phases. There is no evidence for the existence of Zr3Rh5. Based on the results in Rafs. 1 and 5, on the present work on Zr-Rh, and on several additional alloys investigated, the portions between the AB and AB3 stoi-chiometry for Ti-Rh, Zr-Rh, and Hf-Rh are as follows: Further, several ternary alloys near Ti3Rhs and Hf,Rhs were prepared in which it was attempted to replace titanium and hafnium partly by zirconium, niobium, tantalum, and germanium. The results will be discussed in a later section. Structure Determination of Ti3Rh5. Since Ti3Rh5 and Hf3Rh5 are isostructural, the following discussion will largely deal with the former. Although the powder pattern of TisRhs is complex, as found previously,1 it could be indexed by comparison with other structures of A3Bs stoichiometry. Ti3Rh5 was found to be isostructural with Ge3Rh5, whose orthorhombic structure had been elucidated by Geller.9 As both the sizes and atomic numbers of germanium and titanium are comparable, the unit cell volume and the peak intensities could be expected to be similar; however, significant differences exist in the atomic positions, as will be shown. All lines in the powder patterns of Ti3Rh5 and Hf3Rhj could be indexed with primitive orthorhombic unit cells with the lattice constants: The fractional errors are 10 The low-angle portion of the indexed powder pattern of Ti3Rh with sin2 8 < 0.30 is listed in Table I. The extinction laws Okl only with k = 2n and hOl only with h - 2n are compatible with the space group Pbo2 and the more symmetrical space group Phnm of Ge3Rh5. Finally, the positional parameters of Ti3Rh5 and HfsRhs were refined under the assumption that titanium and hafnium occupy the germanium positions in Ge3Rh5. Integrated intensities were obtained from the diffraction patterns by planimetry. Intensities of overlapping reflections were separated by an iteration process incorporated into the least-squares positional refinement program according to a method described previously. The intensities of Ge3Rh5 were used in the first separation cycle, while the atomic parameters of Ge3Rh5 were used as starting values in the first refinement cycle. Absorption due to specimen

Jan 1, 1970

By D. J. McAdam

In a recent survey of the literature, the author has found evidence incompatible with prevalent views regarding the technical cohesive strength and yield strength of metals. Some of the evidence regarding the technical cohesive strength was presented in a previous paper,21 and views were presented that are believed to be in better accord with the evidence. In that paper, it was shown that the technical cohesion limit (contrary to prevalent views) varies with the combination of principal stresses,t and that the influence of plastic deformation on the technical cohesive strength is not in accordance with prevalent views. In this paper, consideration will be given to both the technical cohesion limit and the yield stress as affected by the combination of principal stresses. Evidence will be presented that the variations of the yield stress and the technical cohesion limit with the combination of principal stresses are not in accordance with prevalent views. Consideration will also be given to ductility and work-hardening capacity, which depend on the initial difference between the technical cohesion limit and the yield stress. By the "technical cohesion limit" is meant the mean stress at actual or poten- tial* fracture. An example of a rechnlcal cohesion limit is the "true" breaking stress of a tension-test specimen, the breaking load divided by the sectional area at fracture. The cohesion limit thus obtained for a ductile metal, however, has been greatly altered by the plastic deformation. It is important to consider the cohesion limit of the metal prior to this deformation, and the variation of the cohesive strength with plastic deformation. The relation between the cohesion limit and the corresponding flow stresst determines the ductility of a metal and the energy that it will absorb before fracture. In this paper, as in the previous paper,21 tensile stresses will be viewed as positive and compressive stresses as negative. The algebraically greatest principal stress will be designated S1, the least principal stress will be designated S3, and the intermediate principal stress will be designated Sz In the previous paper, attention was confined to stress combinations with two of the principal stresses equal (with the intermediate principal stress S2 equal eithel to S3 or 31). In this paper, consideration will also be given to the effect of stress combinations with all three principal stresses unequal. In the earlier sections of the paper (pp. 3I I to 324), attention will be confined to stress combinations with

Jan 1, 1942

By D. J. McAdam

In a recent survey of the literature, the author has found evidence incompatible with prevalent views regarding the technical cohesive strength and yield strength of metals. Some of the evidence regarding the technical cohesive strength was presented in a previous paper,21 and views were presented that are believed to be in better accord with the evidence. In that paper, it was shown that the technical cohesion limit (contrary to prevalent views) varies with the combination of principal stresses,t and that the influence of plastic deformation on the technical cohesive strength is not in accordance with prevalent views. In this paper, consideration will be given to both the technical cohesion limit and the yield stress as affected by the combination of principal stresses. Evidence will be presented that the variations of the yield stress and the technical cohesion limit with the combination of principal stresses are not in accordance with prevalent views. Consideration will also be given to ductility and work-hardening capacity, which depend on the initial difference between the technical cohesion limit and the yield stress. By the "technical cohesion limit" is meant the mean stress at actual or poten- tial* fracture. An example of a rechnlcal cohesion limit is the "true" breaking stress of a tension-test specimen, the breaking load divided by the sectional area at fracture. The cohesion limit thus obtained for a ductile metal, however, has been greatly altered by the plastic deformation. It is important to consider the cohesion limit of the metal prior to this deformation, and the variation of the cohesive strength with plastic deformation. The relation between the cohesion limit and the corresponding flow stresst determines the ductility of a metal and the energy that it will absorb before fracture. In this paper, as in the previous paper,21 tensile stresses will be viewed as positive and compressive stresses as negative. The algebraically greatest principal stress will be designated S1, the least principal stress will be designated S3, and the intermediate principal stress will be designated Sz In the previous paper, attention was confined to stress combinations with two of the principal stresses equal (with the intermediate principal stress S2 equal eithel to S3 or 31). In this paper, consideration will also be given to the effect of stress combinations with all three principal stresses unequal. In the earlier sections of the paper (pp. 3I I to 324), attention will be confined to stress combinations with

Jan 1, 1942

By D. J. McAdam

IN a recent survey of the literature, the author has found evidence incompatible with prevalent views regarding the technical cohesive strength and yield strength of metals. Some of the evidence regarding the technical cohesive strength was presented in a previous paper,21 and views were presented that are believed to be in better accord with the evidence. In that paper, it was shown that the technical cohesion limit (contrary to prevalent views) varies with the combination of principal stresses,† and that the influence of plastic deformation on the technical cohesive strength is not in accordance with prevalent views. In this paper, consideration will be given to both the technical cohesion limit and the yield stress as affected by the combination of principal stresses. Evidence will be presented that the variations of the yield stress and the technical cohesion limit with the combination of principal stresses are not in accordance with prevalent views. Consideration will also be given to ductility and work-hardening capacity, which depend on the initial difference between the technical cohesion limit and the yield stress. By the "technical cohesion limit" is meant the mean stress at actual or potential* fracture. An example of a technical cohesion limit is the "true" breaking stress of a tension-test specimen, the breaking load divided by the sectional area at fracture. The cohesion limit thus obtained for a ductile metal, however, has been greatly altered by the plastic deformation. It is important to consider the cohesion limit of the metal prior to this deformation, and the variation of the cohesive strength with plastic deformation. The relation between the cohesion limit and the corresponding flow stress† determines the ductility of a metal and the energy that it will absorb before fracture. In this paper, as in the previous paper,21 tensile stresses will be viewed as positive and compressive stresses as negative. The algebraically greatest principal stress will be designated S1, the least principal stress will be designated S3, and the intermediate principal stress will be designated S2. In the previous paper, attention was confined to stress combinations with two of the principal stresses equal (with the intermediate principal stress S2 equal either to S3 or S1). In this paper, consideration will also be given to the effect of stress combinations with all three principal stresses unequal. In the earlier sections of the paper (pp. 1 to 14), attention will be confined to stress combinations with

Jan 1, 1942

By S. J. Aresco, George E. Keller, J. Visman

INTRODUCTION The accurate sampling of coal, as with most minerals, is a difficult task. Coal is a very heterogeneous material made up of different types of coal and varying amounts of mineral matter. The product as mined may contain all of the layers of coal and impurities found in the seam as well as portions of the strata above and below the coal seam. The preparation plant is the device to size, crush and/or remove impurities so that the coal may be shipped as a saleable product. Sampling may be needed at many stages in the preparation process. Although this chapter is mainly concerned with sampling in connection with preparation processes, the principles will apply to all coal sampling. The main reasons for sampling in connection with coal preparation are: 1. To determine the washability characteristics of the raw coal by means of float and sink tests. These may be used for the design of a preparation plant, prediction of results to be obtained or for monitoring raw coal delivered to the plant. 2. To check the performance of various operating units in the plant, either by float and sink tests or chemical analyses. 3. To determine the analytical characteristics of the products produced, such as moisture, ash, Btu, sulfur, etc. HISTORY The earliest accepted methods of sampling coal in the United States were given in U. S. Bureau of Mines publications by J. A. Holmes,1 G. S. Pope2 and A. C. Fieldner.3 The American Society for Testing and Materials more or less adopted the methods proposed by the U.S. Bureau of Mines in their Standard D-21-1916, and it remained a standard for many years. Briefly, the standard called for a 500 pound sample for coals less than. 3/4 -in. in size and a 1,000 pound sample for all coals over ?4 -in. in size. No mention was made of increments but there was an instruction that the sample should spread over the coal to be sampled. Sample preparation was by hand crushing with the alternate shovel method for reduction to 250 pounds followed by coning and quartering below 250 pounds. Obviously, with the amount of work involved, this method of sample preparation was rarely followed. This method of preparation resulted in samples that were practically air dried because of long exposure while crushing, shoveling and quartering. In the late nineteen twenties and early nineteen thirties, studies for more practical methods were begun. Dr. E. S. Grummell4, Imperial Chemical Industries, Ltd, London, summarized the work to date at the

Jan 1, 1968

By James L. Poole

INTRODUCTION Companies in the mining industry are subjected at times to currency exchange rate risks and interest rate risks. The former occurs any time a firm deals in more than one currency. The latter occurs anytime a firm is financing itself with borrowings that have a floating interest rate (although a fixed rate can be a problem at times). FOREIGN EXCHANGE RISKS There are two general reasons why exchange rate risks (FX risk) can occur: First, a firm may be building a mine or plant in another country where financing for the plan is denominated in funds other than the country hosting the plant. For example, if a Canadian firm is building a smelter in another country which is being financed with Canadian dollars, the C$100 million for labor, which must be paid in the local currency, has subjected the firm to a C$100 million FX exposure risk. If after the cost estimates were made the local currency appreciates 30% against the Canadian dollar, the cost of labor has increased to C$130 million. (Of course, the local currency could also devalue during this period making the labor cost component less than estimated.) The other reason exchange rate risk arises is due to the fact that the costs of production in one country producing minerals may be incurred in one currency while the production is sold in another currency, such as coal being exported from Canada to Japan under contracts denominated in Yen. In this example, the Canadian firm is exporting coal to Japan and being paid in yen in 30 days at a contract price of 100 Y per ton. Furthermore, assume that the mining firm anticipates a depreciation of the yen against the Canadian dollar of 5% in the next 30 days. Therefore, the mining firm stands to lose 5% of the value of its shipment in a month&apos;s time not including carrying costs. In both cases, the mining company risks an adverse change in the exchange rate, that is, the currency the company holds may decrease in value relative to the other currency. There are a number of ways to mitigate the exposure to foreign exchange which can be utilized by the mining company. Optimally, a company should try to match all their inflows and all their outflows in the same currency. For example, if a Canadian firm has mining costs and debt service costs both denominated in Canadian dollars then 100% of their receipts would need to be in Canadian dollars. If a Canadian firm has 70% of its cash outflows in Canadian dollars and 30% of its cash outflow in American dollars in the form of a loan amortization then their receipts would need to be 70% in Canadian dollars and 30% in U.S. dollars. This general principle of matching cash receipts and cash costs in the same currency works fine in a perfect world but what of the real world when these cannot be matched? In actual practice, however, the financial markets have developed a number of ways in which foreign exchange risks can he managed or reduced. FORWARD MARKETS One method is to use the forward markets which are made on a world-wide basis by the commerical banks for foreign exchange. If a firm wanted to lock in the current exchange rate for the receipt of a foreign currency at some point in the future, they could do so by contracting on the forward market with a bank selling that foreign currency at the current rate at the future date. For example, assume that a Canadian firm is exporting coal to Japan which would pay yen in 90 days. If a firm wanted to lock in today&apos;s exchange rate they could contract with a bank to sell on the forward market yen in 90 days. When the yen was received the contract would be executed by selling the yen to the bank and receiving the previously agreed upon number of Canadian dollars. Generally speaking, the forward market can be used to sell forward about 12 months. The costs of selling (or buying) forward

Jan 1, 1985

By Scott Turner

MY YEAR OF SERVICE as president of the A.I.M.E. came at a time when the mineral industry had suffered severely because of disturbed economic conditions throughout the world. The Institute, an integral part of this industry, had been similarly affected. Its membership and revenues decreased; it was confronted by problems of retrenchment at a time when its vigorous operation was especially important to its members, of some moment to the mining industry, and of value to the nation. Throughout the last year, the recession of activity in our industry continued with slight abatement. Your Board accordingly exercised such economy within the Institute as was possible without impairment of essential functions. It endeavored to aid in sustaining the morale of members, their sense of professional unity, and their interest in maintaining research and the interchange of scientific and technologic data upon which the progress of our industry depends. Effort was made to assist in the difficult task of finding employment for needy members; our endeavors have met with some success.

Jan 1, 1933