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By Caudle B. H., Cooper H. E. Jr., O. K. Kimbler

The results of studies of areal sweepout behavior in nine-spot injection patterns are presented in this paper. The ratio of the production rates of the two types of producing wells in a nine-spot pattern is considered in its effect on breakthrough. A model study was used to develop a method to predict the production histories of both side and corner wells and to develop the production history of the pattern or the field. INTRODUCTION The importance of areal sweepout behavior in injection operations has been recognized for a number of years. We now know that the fraction of a reservoir which will be reached by an injected fluid is a function of the geometric arrangement of the production and injection wells and of the mobility ratio of the fluids. Of these two, the mobility ratio (which is a measure of the relative ease with which the oil and the injected fluid can move through the reservoir) appears to have the greatest effect on recovery. The effect of mobility ratio on production history has been reported for most of the dispersed injection patterns (i. e., five-spot, staggered and straight-line drives).4-6 ere has been very little published on the nine-spot injection pattern. This may be because the two types of producing wells in the nine-spot pattern can—and sometimes should —be produced at different rates. This adds a second variable (the fist being mobility ratio) to be covered in a study of the nine-spot flooding network. Muskat1 has reported the sweep pattern efficiency at breakthrough for a mobility ratio of one at various ratios of producing rates for the two types of wells. For these conditions he found maximum oil production at breakthrough to occur when the corner well (the producing well farthest from the injection well) is produced 10.7 times as fast as the side well. Other investigatorsa2,3 have studied a few of the many possible operating conditions possible with this injection pattern. This paper presents the results of studies which are strictly applicable to a nine-spot injection pattern surrounded by other nine-spots in a homogeneous reservoir and which has no flowable gas present (no oil bank is formed). Although the full ranges of the variables mentioned above have not been covered, the results reported here should allow reasonable estimates of field operations to be made. EXPERIMENTAL The well spacing geometry of the nine-spot pattern is shown in Fig. 1. Here we see that each injection well is surrounded by eight production wells arranged in a square. Four of these producing wells are located at the corners of the square; the other four are in the midpoints of the four sides. These two types of producing wells are called "comer wells" and "side wells" respectively. It is the ratio of the producing rates of these two types of wells that affects the breakthrough recovery, as calculated by Muskat.' This ratio is defined as the total producing rate of a corner well divided by the total producing rate of a side well. The model used in this study represented one-eighth of a nine-spot pattern, as shown in Fig. 1. It contained the equivalent of one-fourth of a side well, one-eighth of a comer well, and one-eighth of an injection well. Because of these proportions, the actual producing rate ratios in the model were one-half of the ratios of producing rates as

Jan 1, 1965

By Olaf N. Rove

YEAR 1951 has been bright for the mining geologist. He has arrived after struggling for a generation or two to sell his wares through service to the operator, the mine superintendent, and the manager. But as he establishes himself on the operating front, he is faced with even greater challenges; namely, the development of basic geologic principles governing ore-forming processes and the occurrence and distribution of mineral deposits. These principles are needed to guide his ever-widening search into new areas. He must now stimulate the interest, confidence, and imagination of the mining executive and layman in fostering the research essential to expand rapidly the basic science of geology and its more specific applications to the mining industry. The hallowed time-tested tools of the prospector and the industry are no longer adequate. New basic data, principles, and ideas are needed to guide industry's exploration. The entry of Government into exploration, by lending money to prospectors and operating companies to partly finance exploration, is one of the major events of the year. This was one of the activities for which the Defense Minerals Administration was established in the Department of the Interior. DMA was staffed largely with men from industry, but the Bureau of Mines and the Geological Survey have acted as its technical advisors and have provided its field staff. Since the start of this work in May 1951 to November 19, 1951, the DMA has approved 208 exploration projects while denying 318 others. The contracts involve about eleven million dollars worth of exploration, of which Government's share is about sixty percent. A few contractors have already made discoveries and are into production to repay their loans; others have completed their projects without making a discovery. In November, the DMA was split into two successor units. The Defense Minerals Exploration Administration was set up in the Interior Department to carry out the exploration program. The Defense Materials Procurement Agency was organized as a separate agency to carry out the other functions of the former DMA, as well as general strategic materials procurement. The strategic materials group of the Economic Cooperation Administration was transferred to DMPA in order to centralize the procurement from foreign sources into one agency, although the authority and scope of activities of this group have not been announced as of the middle of December. The DMA-DMEA-DMPA programs are making large demands upon the technical manpower of the Geological Survey and Bureau of Mines. It is still too early to evaluate with confidence the productiveness of this diversion so far as the immediate and long-range effects are concerned. Many of the individuals associated with these programs, as well as others in the industry, are convinced that from a long-range national viewpoint, the regular programs of both the Survey and the Bureau equal or even exceed in importance the current emergency mineral activities of these two agencies. Uranium Uranium continued to be the glamour-girl metal of the mineral industry, and at the request of the Atomic Energy Commission uranium and thorium were included in the DMA exploration program. Industry's intensive prospecting and mining activities in 1950 were further expanded early in 1951 when the Atomic Energy Commission announced an increased price schedule and a new bonus payment

Jan 1, 1952

By G. W. Nabor, H. J. Ramey

A blotter-type electrolytic model was utilized to prepare flow diagrams for a field test of the in-situ combustion process. It is pointed out that the areal sweep of a combustion pattern is similar to sweep patterns that would be developed at an infinite mobility ratio, which exists (approximately) across a combustion front because of the complete removal of liquids from the sand behind it. The precision of the blotter method was tested by comparison with results obtained by other techniques and was found to be satisfactory. The blotter-type model will not furnish as much information as more elaborate and expensive potentiometric models, but its speed of operation and ease of construction make it a highly satisfactory tool to determine areal sweep patterns. A tabulation of sweep efficiency and mobility ratio is furnished for various well geometries. INTRODUCTION In connection with the first field experiment with the in-situ combustion method of oil recovery," flow diagrams for the unique well geometry and boundary conditions involved in the subject test were developed. From these diagrams, information concerning the areal sweep and distribution of flow was obtained. A survev of the literature on potentiometric models, utilized to solve two-dimensional flow problems,1,2,3,4,5,6,8,9,10 indicated that the information desired could be obtained by means of a blotter-type model by making certain modifications in the technique. The following discussion describes the application of the modified blotter-type electrolytic model to the solution of the problem of the flow of gases in an oil reservoir for an engineering appraisal of the progress of the combustion front during the first field experiment. Considerable information has been published on the theoretical aspects of electrical model techniques. One author stated that:' "The models are based on the observation that since the velocity of an ion in an electrolytic system is proportional to the potential gradient, just as the velocity of a fluid particle in a porous medium is proportional to the pressure gradient, the paths of the ions in the electrolytic systems must be similar to those of the fluid particles in a porous medium with the same geometry and equivalent boundary conditions." This suggestion of a real analogy has been established repeatedly by comparisons of analytic derivations with model results. The analogy between the movement of a combustion front and the model results is not as apparent, since the combustion process differs from water flooding and gas cycling in that the velocity of the burning front is different from the velocity of the injected fluid at the front in combustion. Analogy between electric current flow and movement of the burning front exists because the local velocity of the burning front is proportional to the air (mass) velocity at this point (assuming frontal temperatures are at least as high as the ignition temperature of the fuel). Most model studies fall into one of two categories: (1) electron conduction, or (2) ionic conduction models. Examples of these are models which employ as flow media either a thin metal sheet, or a porous blotter saturated with an electrolyte, respectively. Of the two, the porous blotter model, an ionic conduction model, appears to be the most adaptable to the flow problem that results when using in-situ combustion. This problem requires that the mobility of the particles in the model be controlled analogous to the mobility of gas flowing through burned sand of high permeability to gas and then through oil sand of lower permeability to gas. Although the blotter model has heretofore been

Jan 1, 1955

By M. R. Annis

Drilling mud rheological and gel property changes due to elevated temperatures frequently cause problems in drilling deep wells. A laboratory investigation of the effects of time, temperatures up to 3001; and mud composition on these properties was made with concentric-cylinder, rotational viscometers of the Farm type. The viscometers were sealed from the atmosphere to prevent dehydration of the sample. Gel strengths were found to be more sensitive than viscosities to changes in temperature and mud cornposition. The high-temperature gel strengths of bentonitic clay drilling fluids were not predictable from low-temper.atl4t-e measurements. At high temperatures, [lie clay flocculates. The floccu-lation temperature depends on the concentration and type of solids in the mud and the degree of chemical defloccula-tion. As the bentonite content of a mud increases, floccrrla-tion temperature decreases. Lignosulfonnte treatment raises the flocculation temperature. To minimize adverse high-temperatrrre effects on mud properties, the mud should be maintainted at minimum solids content, properly treated with. a defloccullant which is stable at the maximum ten?perature encountered and maintained at a pH high enough to insure activation of the deflocculant for the duration required. The Methylene Blue Test for cation exchange cnpacity of a clay was found to be a useful guide in deter.niinirig when a mud contains a dangerously high clay conWnt. INTRODUCTION In drilling a well it is desirable, for several reasons, to know the effects of time and temperature on the rheological and gel properties of the drilling fluid in the hole. Problems associated with drilling fluids normally occur down-hole: so it is reasonable to be interested in the properties of drilling fluids at the conditions which exist down-hole. It has long been observed that high pump pressures are often needed to start circulation after a trip and that the "bottom-up" mud that is circulated back to the surface is usually more viscous than it was before "cooking" at the bottom of the hole for several hours. Burk-hardt' showed that a knowledge of rheological proper-ties of muds under down-hole conditions is necessary to predict pressure surges. He showed that the viscosity of a mud greatly influences the magnitude of pressure surges and the likelihood of lost returns. McLean et aL~ showed that the rheological properties of a mud under down-hole conditions are very important to cementing operations. This paper presents data showing the effects of time and temperature on rheological and gel properties of water-base drilling fluids. Measurements of gel strengths and rheological properties were made with two different concentric-cylinder rotational viscometers of the Fann type. One was the conventional six-speed viscometer modified to operate at temperatures up to 350F and at the vapor pressure of the liquid at test temperature. The other viscometer was newly designed to operate at temperatures up to 500F and pressures up to 1,000 psi. Tests with both instruments were made under essentially the same operating conditions. The new viscometer is shown in Fig. I. In this investigation, great emphasis has been placed on the rheological properties of clay suspensions, since clays are largely responsible for a mud's thixotropic and gel-

By Vincente M. Jr. Lim, Brain W. Lawrence

Underground mining operating costs consist pri¬marily of direct costs such as labor, supervision, ma¬terials, and supplies, plus indirect costs. The material and supply costs vary substantially between mines but commonly range from 35 to 60% of the total direct costs. As part of the preparation of accurate operating cost estimates for a project, information must be as¬sembled to select the material and supply items and to determine the quantities of these items that will be required. Cost data then can be incorporated with the estimated material and supply requirements to determine the total material and supply costs for the project. Estimated quantities will vary considerably between mining projects depending upon factors such as the type of material being mined, the mining method, the equip¬ment selected, drilling and blasting techniques, mucking and haulage methods, ground support requirements, number of men employed, etc. Valuable information often can be obtained from suppliers and manufacturers, including data on current operating practices for similar projects, recommenda¬tions for the specific project, and preliminary cost quota¬tions. However, quotations obtained for estimates are generally higher than the prices paid at the time of purchase. Mines operating in the area, or those mining under similar conditions, also can provide essential design information, but it often is difficult to obtain cost information from the operators. Estimated costs for the most common major items of underground materials and supplies are presented in the accompanying tables. They represent current costs or cost ranges and are intended to be used only as a guide when preparing budgetary estimates for materials and supplies. Cost data were obtained mainly from suppliers and manufacturers, with up to three quotations being ob¬tained for each item whenever possible. In most cases, as noted in the table footnotes, prices were based on quantities representing normal requirements for me¬dium-size mine operations. Every attempt was made to assure accuracy. However, the cost for most items, especially those with multiple sources, can deviate sig¬nificantly for various reasons such as geographical loca¬tion, current market conditions, quantities ordered, size of the customer (which could influence the ultimate quantities purchased), and the ability of the purchasing department. It is also for these reasons that the actual amount individual mines are now paying for specific items may vary from the estimated cost in the tables. The cost for specific sizes of items that do not appear in the tables often can be factored from the data pre¬sented. Costs for many items can be converted easily to a unit that sometimes is more basic, but not as com¬monly used, such as from a length to a weight unit. Other items not listed, such as repair parts for equip¬ment, will involve contact with suppliers to determine specific recommended requirements and to obtain costs for these items as related to each individual project. Costs are expressed in terms of US dollars as of 1978 unless otherwise specified. All measurements, sizes, and weights are expressed in conventional units (i.e. feet, inches, pounds) ; applicable metric conversion fac¬tors follow each table. Costs are represented for drill steel and bits; mining explosives; blasting caps; blasting supplies; ground sup¬port items; ventilation equipment; utility pipe, utility hose, and accessories; mine track and accessories; tires for underground equipment; lubricants and fuels; and miscellaneous supplies.

Jan 1, 1982

By S. M. Brisbane

THE life of refractory linings in rotary cement kilns can be shortened by uneven or inadequate support of the kiln and its shell and by variations in the conditions of operation. Alteration of these factors may, according to their nature, bring about mechanical deterioration of the lining or promote chemical attack. Chemical attack has been described elsewhere.1-4 The distortion of cement kilns has been discussed in a general manner by Rochow,6 ho indicates the amount of distortion which may occur in an unlined shell, and by Rosenblad: who states that the deformations of the shell are greatest above the trunnions. The present article indicates the extent of the distortion of a kiln to an oval shape in certain selected positions and the deviation from a straight line of the kiln along its length between two supports. The cement kiln is a long, hollow, rotating cylinder consisting of steel plates which may be either welded or riveted together. This shell is supported on tires, frequently 60 ft apart, which roll on trunnions. The large mass of the kiln causes distortion of its cross-section; it becomes slightly oval, i.e., diameters lengthen as they become horizontal and shorten as they become vertical. There is also a tendency for the shell to bow between supporting tires; this bowing may be upwards or downwards depending on the position of the weight forces and the supports. The extent of the deviations from the ideal kiln shape was investigated. Observations described in this paper were made on the shell of the clinkering zone between the first and second tires of two of the kilns at the Geelong works of Australian Portland Cement Pty. Ltd. The kilns were both 400 ft long and 10 ft diam with a welded enlarged burning zone 11 ft 6 in. diam. The shell was fabricated from mild steel and in this section its thickness was 1 in. It was lined with 41/2 in. of chemically bonded magnesite-chrome bricks which were jointed with mild steel plates. The kilns operate on a 4 pct slope and the speed may be varied between 0.75 and 1.68 rpm. The wet process is used; kilns are coal fired and are fitted with integral coolers. Temperatures within the kiln are difficult to determine with any accuracy because of the absence of black body conditions, but surface temperatures are probably in the range 1420° to 1470°C. Average shell .temperatures at the two positions at which the deformation studies were made were 370°C and 355"C respectively. Methods: In determining the ellipticity it was necessary to measure the lengths of the vertical and horizontal diameters while the kiln was in operation; data obtained from a stationary kiln would not have the same practical significance. For this purpose the four points V, W, X, and Y at the ends of two mutually perpendicular diameters were marked on one kiln. The point X is shown in Fig. 1. For the measurement of the length of the vertical diameter a vertical column, to which a horizontally aligned telescope was attached, was placed 10 ft from the kiln. The telescope could be racked up and down in a vertical plane. Two sets of readings were taken, one set when each of the points V, W, X, and Y was at its maximum height above the

Jan 1, 1957

By W. W. Gerberich

SINCE the original formulation of fracture mechanics concepts,&apos;y2 relatively little work has related the fracture mechanics description of metal behavior on a gross, structural component scale to the behavior of its microstructural constituents. The purpose of this note is to suggest the utilization of fracture mechanics to describe material behavior on a microscopic scale. A Fracture Concept for Two-Phase Structures. A critical displacement concept considering the tough unidirectional phase in a relatively brittle matrix is first examined. It is assumed that a flaw exists at position 1 as depicted in Fig. 1. As the crack moves from position 1 to 2 the plastic zone could be of the shape described with the size approximated by the Dugdale3 model: where c is the half-crack length, OO is the applied stress, and oyS is the yield strength. Hahn and Rosen-field4 have presented a simple fracture criterion for homogeneous materials based upon the displacement at the crack tip. They also have shown the equivalence of the critical displacement, v*, to fracture toughness by Kc = (2v*oysE)1/2 [2] The displacement at the crack tip is given by and the displacement at any distance in front of the crack tip may be determined knowing the plastic zone size and utilizing the development of Goodier and Field.5 Recalling Fig. 1, the value of v* for fracture in material M1 may be calculated from Eq. [2] providing the critical stress intensity factor is known. As the material is loaded, when vc > vM1 the crack moves from position 1 to position 2 where it encounters the tough second phase. If vc2 < vm2, then the crack will be arrested providing the width of the M2 phase is sufficiently large. This condition is necessary since the displacement at position 2 has increased due to the longer crack length. Thus, the displacement at the distance X (representing the dimension of the second phase) in front of the crack may be larger than the critical displacement for failure in the brittle material. If v, > vm, then failure will proceed in the brittle material. The value of v, may be calculated as indicated above. This latter consideration is the same as predicting fracture of a brittle second phase at some distance in front of a crack located in a ductile matrix. Experimental Evidence. On a macrostructural scale, tungsten wires in a ductile Be-Cu matrix are seen to fracture well ahead of the main crack front in Fig. 2. Utilizing a Kc value of 14.2 ksi-in.1/2 obtained by Schroder et al.6 for as-received tungsten, the critical displacement 2v*, for fracture is calculated from Eq. [2] to be about 6.5 X 10-5 in. The crack tip displacement, 2vc, is calculated from Eq. [3] to be 0.0028 in. indicating that fracture would occur at a considerable distance in front of the crack tip. From the ratio of v*/vc and the plastic zone estimate of Eq. [I], the distance in front of the crack tip at which a tungsten wire should fracture is calculated to be about 0.195 in. As shown in Fig. 2, a fractured wire occurs at a distance of 0.140 in. (fractured in three places), not at 0.250

Jan 1, 1968

By J. L. Chakravatti

The movement of air from the surface, through the workings and then back to the surface, is the most important consideration in the ventilation design. The purpose of mine ventilation is seen as the maintenance of acceptable atmospheric environments wherever men work or travel underground; and acceptability is related to the risk of harmful biological responses associated with exposure to atmospheric environments in mines. In mine ventilation, a design engineer often has to contend with a number of problems not experienced to the same degree in other industrial situations; i.e. the vast extent of many workings - offering continuously moving work locations; the difficulty of predicting. inflows of contaminants to the airways which contain the micro-environment; and especially, the effect of mining activity within the duct system (e.g. mine roadways) carrying the airflow. Mining of uranium does not differ, in many respects, from the mining of other metals; the ore must be drilled, blasted and transported from the mine. Underground mining proceeds by the continous development and advancement of a network of passages and rooms or stopes from which ore is excavated. As in most mines, rock conditions and ore grades in uranium deposits very from area to area and between different mines. 1 In uranium mining, in addition to eliminating noxious gases from blasting, diesel equipment exhaust, and the control of dust, it is also necessary to dilute and remove radon, the gaseous element in the uranium chain and its decay products. Long term epidemiologic studies of United States uranium miners and that of miners in Czechoslovakia have confirmed that there is an association between excessive ex¬posure to radon daughters and an increased risk of lung cancer. The common practices for controlling these airborne contaminants are: (1) the circulation of large volumes of fresh air from surface to dilute these contaminants to a permissible concentration, and (2) the sealing of worked-out areas to prevent ingress of radon into ventilating air. 2 The sealing of worked-out areas can be easily and effectively accomplished, but the contaminants intro¬duced from the workings may raise the concentrations in the airways despite a good ventilation system and use of the maximum available amount of fresh air. The ventilating air can be strategically filtered to reduce dust and radon daughters to acceptable levels, and reused prior to final discharge. This strategy properly applied, will decrease the air colume requirements in certain marginal areas of the mine and thereby avoid an overall increase in the mine airflow. This is particularly advantageous in the cold climates of Canada, where the incoming fresh air must be heated in the winter months to a comfortable temperature, Thus, an economical and technically feasible air filtration system, strategically placed, can improve the underground mining environment and reduce the overall cost of ventilating the mines. In addition to affording flexibilty to the venti¬lation system, a potential saving of $1,500/m3/s/year would be realized if air is successfully cleaned and reused when compared with the conventional methods of generating an air current. 3 Denison is justly proud of its achievements in the air filtration technology development. This paper limits itself to an examination of the radiation control aspect of the air filtration strategy.

Jan 1, 1980

By J. B. Clark, A. S. Yue

PfANN1 has shown that when a cylinder of molten binary alloy freezes directionally, the solute distribution can be described analytically by cg=k(l-g)k-1 [1] in which Cg is the solute concentration in the solid at the point where a fraction g of the original liquid has frozen and Co is the initial solute concentration. The term k is the effective distribution coefficient of the solute and is defined as the ratio Cx/CL where Cl is the solute concentration in the main body of the liquid. Both the slope and the intercept at -log (1 - g) = 0 of a logarithmic plot of Eq. [I.] provide a measure of k. Under conditions of freezing where diffusion in the solid is negligible and the solute concentration in the liquid is uniform, the effective distribution coefficient k, derived as above, is equal to the equilibrium distribution coefficient taken from the phase diagram. However, under ordinary freezing conditions, even though solid-state diffusion is negligible, the solute rejected from the freezing solid does not have sufficient time to diffuse uniformly throughout the main body of the liquid. A concentration gradient is formed in a boundary layer in front of the advancing solid-liquid interface as pointed out by Rutter and Chalmers.&apos; Thus, under these conditions, the concentration of the freezing liquid in the boundary layer is higher than that predicted from the equilibrium phase diagram and consequently the effective distribution coefficient k, obtained from directional freezing experiments, is larger than the equilibrium distribution coefficient. The symbol k will designate the effective distribution coefficient in the discussion which follows. In this paper, the determination of the effective distribution coefficient is demonstrated for several dilute binary alloys and a polynary magnesium alloy. The effect of freezing rate on the value of the derived k is presented in detail. An example for the calculation of the solidus in a dilute binary solution using the experimentally determined effective distribution coefficient is given. MATERIALS STUDIED Binary magnesium alloys of copper, nickel, and zinc containing approximately 0.10 wt pct solute and a polynary alloy containing approximately 0.10 wt pct each of several elements were selected for study. The spectrographic analyses of these alloys are given in Table I. The analyses are believed accurate to within +5 pct of the stated value. EXPERIMENTAL PROCEDURE Extruded rods of the alloys, 3/8 in. in diam and 11 in. long were used for directional freezing studies. The alloy rods were melted horizontally in a high-purity graphite boat by a Nichrome-wound tube furnace which moved parallel to the length of the rod. The boat was supported inside a Vycor tube suitably connected to provide a protective atmosphere of dry SO2. For an experimental run, the following procedure was used. The alloy rod was cleaned in dilute HNO, and dried. The graphite boat containing the alloy rod was placed within the Vycor tube and the tube connected to the gas train&apos; The &apos;ystem was purged for 20 min with dry argon and then a flow of dried SO, was maintained for the duration of the run. The movable furnace was positioned and either all or a desired portion of the alloy rod was melted. After attainment of steady-state temperature gradients the molten bar was solidified from one end to the other at a controlled rate by pulling the furnace along the length of the bar at the desired speed. At the conclusion of the run, the directionally frozen alloy rod was analyzed spectrographically at 1 1/4-in. intervals.

Jan 1, 1961

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.&apos; 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

Objective As mines become safer and disasters fewer, a time may come when no employees in the mining industry have any practical experience in managing a mine emergency. To fill an emerging gap in mine emergency response training, a multidisciplinary team of researchers and computer programmers from the NIOSH Pittsburgh Research Laboratory has developed an intervention for mining personnel who would be in leadership positions in the command center during emergencies. It is called MERITS, a computer-based emergency simulation exercise. MERITS is short for "Mine Emergency Response Interactive Training Simulation." It provides trainees an opportunity to gain command center experience during a simulated underground coal mine emergency. MERITS allows trainees to practice information gathering, situation assessment, decision-making, and coordination skills without risk to personnel or property. Individuals representing mine emergency response personnel, various State safety officials, and private mine management from three States have participated in MERITS training. What Is MERITS? MERITS is an extension of previous work that compiled a knowledge base relevant to mine disaster management via interviews, analysis of literature, reviews of past events, and observations of mock mine disasters. Findings from prior research found a need for an alternative method of command center training. This new computer simulation is being used to train emergency command center personnel to effectively manage an underground mine fire. Modeled after training simulations developed for other industries (e.g., to address nuclear accidents, chemical spills, evacuations, etc.), MERITS simulates underground and surface events related to the disaster. It exposes the user to events that typically occur during a mine emergency, such as lack of information and miscommunication. It also presents trainees with issues that must be addressed, such as making provisions for briefing news media and victims&apos; families, ordering needed supplies, interfacing with Government enforcement agencies, and housing mine rescue teams. The outcome of the scenario is determined by the users&apos; decisions and their emergency response plans. Field Results Field testing of MERITS was conducted with eight full-day training sessions at four different locations in Colorado, Pennsylvania, and West Virginia. These field tests benefited from the knowledge and experience of State of Pennsylvania inspection officials, representatives of the United Mine Workers of America(UMWA), and private mine management who served as trainees for the MERITS training sessions. Trainees self-reported an average of20 years of work experience in the mining industry. All trainees had participated in Mine Emergency Response Development (MERD) training exercises, and many had been involved in real mine emergencies. Trainee evaluation of the MERITS program is summarized below: •Twenty-six of the 27 trainees believed that the MERITS simulation helped them know how to prepare for a real emergency. •Twenty-five of the 27 trainees believed that the storyline for the MERITS simulation was either "very realistic" or "realistic." •All 27 trainees believed that the MERITS simulation helped them learn how to better handle a real emergency. •None of the 27 trainees stated that they were bored during the simulation. After the initial field tests, MERITS was refined and retested during training sessions for individuals from State agencies, the UMWA, and private industry. Mining personnel from underground coal and underground stone mines participated and responded positively to the learning experience. Technology Transfer Industry introduction to the MERITS program will occur throughout 2002. MERITS presentations at the Joseph A. Holmes Safety Association National Meeting in Virginia Beach, VA, in June.

Jun 1, 2002

By Henry A. Stiff, and Lawrence E. Davis

The authors previously presented a method for predicting the tendency of oil field waters to deposit calcium sulfate. The present paper gives a similar method for calcium carbonate. Methods for predicting calcium carbonate scaling tendencies in fresh waters have been available for some time, but these could not be used for brines. By experimentally deriving the value of the K term in the Langelier equation, a method has been developed which applies to waters of high salt content. A statistical study is included which shows that the experimentally derived values of K are in good agreement with actual conditions. Several applications of the final equation to production practice are given. INTRODUCTION The authors previously presented a method for predicting the formation of calcium sulfate scale in oil field waters.&apos; Although calcium sulfate deposition is important in production operations, the majority of scale problems involve calcium carbonate. The present communication discusses a method by which the formation of this type of scale in oil field waters can be predicted. Calcium carbonate precipitation is caused by a shift toward carbonate in the carbonate-bicarbonate-carbon dioxide equilibrium. When equilibrium shifts in the other direction, the precipitate goes back into solution. Since there is usually considerable delay between the establishment of an equilibrium and the precipitation or solution of calcium carbonate, unstable conditions exist in which a water will precipitate or dissolve calcium carbonate on standing. Tillmans? who did a major part of the early work on carbonate scaling, pointed out that the condition of equilibrium not only indicates the tendency of a water to scale but also is an indication of its corrosive properties. Previously precipitated calcium carbonate combines with iron to form a dense crust which inhibits corrosion. If the water tends to dissolve carbonate, the scale becomes porous, and electrolytic corrosion takes place. Tillmans1 work was extended by several investigators until in 1934 Langeliers developed an equation setting forth the conditions of carbonate equilibrium. By the use of this equation the pH of a water at equilibrium can be calculated. If the actual pH is higher than the calculated pH, the water has a tendency to form scale. If it is lower. the water has a tendency to be corrosive. Langelier&apos;s equation can be expressed in a simple form as follows: SI = pH-pCa-pA lk-K where: SI is the stability index. A positive index indicates scale formation. A negative index indicates corrosion. pH is the pH of the water sample, as actually determined. pCa is the negative logarithm of the calcium concentration. pAlk is the negative logarithm of the total alkalinity. K is a constant, the value of which depends on the total salt concentration and the temperature. This equation has been shown to apply to waters with totaI solid concentrations as high as 4.000 ppm. In fact, control of fresh water treatment by means of this equation has been standard text-book practice for a number of years. Nomographs have been worked out so that the stability index of a fresh water sample can be determined in a matter of minutes. Most oil field waters, however, contain well over 4,000 ppm of salts and for this reason the usual application of Langelier&apos;s equation can not be made. By an empirical method we have extended the application of the Langelier equation to waters of high salt concentration. By the use of this equation the tendency of oil field waters to deposit calcium carbonate can be predicted.

Jan 1, 1952

By D. W. Loughridge

INTRODUCTION Raising capital in the 1990s is a timely subject. However, many of us thought why should raising capital in the 1990s prove to be unlike it was in the 19805, the 1970s, the 1%0s, or earlier? There are a lot of creative people in banking and in corporate treasury departments. New tools and techniques are continuing to emerge, many of which are discussed in other papers in this book. While we have new financing tools at our disposal, the underlying issue has not changed. Projects and investments must be judged on their own merits. If they cannot be justified by conventional financial analysis, they should not be built. If they can be justified, it is likely that the financing will be available. The challenges faced in raising capital in this decade will persist to the next. The issues of the 1980s -- currency fluctuation, disruptive merger and acquisition activity, commodity price volatility, political risk, and debt crisis -- will all persist in an increasingly international business environment. I will focus on the financing we have just completed for the Escondida Copper Project in Chile. The Escondida financing is structured along the classic project financing model developed in the late 1970s, but it deals with issues of the 1980s and the 1990s. The financing was intriguing, not just because of the technical issues involved, but also the challenge of reaching an agreement in a large and diverse group. Escondida had the one essential feature that made its financing possible. It was commercially very attractive. It would produce at low cost a commodity (copper concentrates) that direct consumers felt would be in tight supply. The project had the direct support not only of its sponsors, but also its customers, and, very importantly, its host government. The non- recourse financing of Escondida would not have been possible without the support of these parties, and without the robust economics that made this effort worthwhile. The issue was more complicated. Operations in all parts of the world entail some political exposure. This project is in Chile, where there was a constitutional requirement that a plebescite be held when the project would be in its early stages. This plebescite would determine the fate of a 15 year old military dictatorship. The project is copper, a commodity generally deemed to be in surplus. Copper has a history of price volatility (whose price has varied from $.58 to $1.53 per pound over the past two years). It is politically sensitive, not only in Chile, but in the United States, Canada and other developed countries. The project is international, with all the problems of dealing with a variety of currencies, a variety of cultures, and a variety of laws and commercial practices. The project is also large and capital intensive, with a total cost of U.S. $1.1 billion. ESCONDIDA PROJECT Escondida, the largest known copper ore body in the world, was discovered in March of 1981, by Utah International in a joint venture with Getty Oil Company. In 1984, both Utah and Getty were involved in ownership changes, Utah being acquired by BHP. A later restructuring of the ownership of Escondida resulted in BHP, as manager, carrying a 60% interest, Rio Tinto-Zinc Corporation (RTZ) 30%, and a consortium of Japanese companies led by Mitsubishi Corporation with 10%. The Japanese consortium is incorporated as JECO (Japan Escondida Corporation). This ownership structure, which spread ownership around the world, was the first tangible step toward managing both the commercial and political risk associated with the project.

Jan 1, 1990

By Robert C. Collins, Clifford L. Jewett

The development and production of artificially colored roofing granules have been key factors in the success of the asphalt roofing industry. They provide protection for the asphalt from solar radiation and satisfy the home-owner&apos;s desire for a wide range of attractive colors. As a result, the asphalt shingle has become the most commonly used roofing material today. In addition to providing a stimulus to the asphalt roofing industry, artificial methods of coloring granules completely changed the nature of granule production from one of quarrying, crushing, and screening to a complex technical operation. Today&apos;s roofing granules are the result of many years of research, and they are produced in modern plants to exact specifications. History Asphalt roofing materials available today are the product of a long evolution of technological development. Forerunners of the present bituminous products, appearing as early as 1780, were primitive constructions of burlap saturated with crude tar or pitch. An early improvement to this basic effort was the dusting of talc, mica, or other finely ground material on the coated surfaces to keep the roofing from sticking to itself in storage. At times some pigmentation was added with the dusting material for initial color, but this was far from satisfactory. In fact, this early product possessed none of the qualifications of a good roof surfacing material. It was not economical; it was difficult to handle and apply; it was not attractive; it provided no resistance to fire; and it would not withstand the elements as well as other products then in use. Methods and machines were soon devised to saturate endless webs of felt with asphalt from the fast-growing petroleum industry. Even though saturated felt became available, this product of rags and wood fiber had almost the same limitations as burlap and tar. Even with heavier coatings of asphalt applied to both sides of the sheet for rigidity and longer life, the product still lacked color and fire resistance. Further, it was found that constant exposure to the sun soon caused the asphalt to become brittle, and it eventually cracked and leaked. This led to the first important use for crushed mineral products on asphalt-coated shingles and provided the key to the subsequent growth of the asphalt roofing industry. When a layer of crushed mineral was embedded in the surface exposed to the weather, the asphalt coating was protected from ultraviolet light and retained its elasticity over a much longer period of time. In addition to this protection, the mineral surfacing provided some resistance to fire. Fire resistance in the past had been found only in the expensive slate or tile roofs and the unattractive and expensive metal roof coverings which were then available. The most common mineral granule applied to the early asphalt roofing products as they began to make their first significant advances after 1900 was screened waste from slate quarries. Slate granules being available in dull shades of red, green, black, purple and a few other miscellaneous colors did provide a degree of color. Since slate has a flat platelike cleavage, it provides a cheap protective cover for the asphalt-coated sheets. Records show that coarse slate screenings were used for asphalt roofing as early as 1906. As the asphalt roofing industry developed, many other natural colored rocks and minerals were used to coat the surfaces of the roofing. The list included silica

Jan 1, 1975

Surface mining of coal involves removing soil and rock that overlay coal beds in order to expose the coal. Compared to underground coal mining, surface mining generally costs less, is safer for miners, and usually results in more complete recovery of the coal. However, it also results in much more extensive, though temporary, disturbance of the land surface, which can cause serious environmental problems unless the mined land is carefully reclaimed. The earliest commercial U.S. coal production was in the anthracite region of Pennsylvania. Beginning in 1807, mining became important enough for the federal government to begin keeping statistics on annual production. During those early years of coal mining, reclamation of mine sites and Support facilities was not required by law; however, in the 1930&apos;s, as surface mining became more widespread, the idea of mandatory environmental protection and reclamation started to grow. In 1939 West Virginia enacted the first law to regulate the coal mining industry. Similar laws were enacted by Indiana in 1941, Illinois in 1943, and Pennsylvania in 1945. During that period, coal production accelerated greatly and the surface mining of coal became much more prevalent. During World War 11, the demand for coal took priority over environmental concerns, with little consideration for reclaiming the land after mining. Mining pits were not backfilled. Dangerous highwalls were left exposed. Trees and other vegetation were buried by waste material that was simply dumped down the slopes below mines. Topsoil was buried or allowed to wash away. Landslides formed on unstable hillsides. Slopes eroded rapidly because of the lack of vegetation. Polluted water collected in mine pits. Streams became clogged with sediment. Streams and rivers were frequently polluted by acid mine drainage. This vivid image of surface coal mining remains indelible in the minds of most Americans, who remain unfamiliar with modem surface coal mining methods and the changes brought about by the Surface Mining Control and Reclamation Act. After 1945, more coal-producing states instituted regulatory pro- grams. Many required permits for mining. Some included inspections. A few made operators post performance bonds to guarantee reclamation when mining was completed. However, due to the variations in those individual programs, it cost less to produce coal in states with less stringent reclamation requirements, giving operators in those states an economic advantage, generally at the expense of environmental quality. With the 1970&apos;s came a surge in the demand for coal resulting from the nation&apos;s need for increasing electric power generation and the uncertain supply and increased cost of imported petroleum-based fuels. The resulting increased coal production prompted a series of state laws, primarily in the West, aimed at controlling the environ- mental effects of surface mining. During the late 1960&apos;s and early 1970&apos;s, North Dakota, Wyoming, Colorado, and Montana passed laws to regulate mining and reclamation. The continuing lack of uniformity among the state surface mining programs and the increase in unreclaimed land and associated pollution of water and other resources caused a growing demand for nationwide regulation of surface coal mining, both to remove the economic inequities and to is sure an acceptable standard of reclamation. After many years of debate, Congress passed the Surface Mining Control and Reclamation Act (SMCRA), which was signed into law August 3, 1977. The Act established a coordinated effort between the states and the federal government to prevent the abuses that had characterized surface coal mining in the past. The law also applied to surface disturbances created by underground coal mining.

Jan 1, 1992

By S. A. Braley

THE first commercial production of bituminous coal in the United States was in 1820, and formation of acid in the areas from which the coal was removed began at that time. Thus it is 130 years since the development of mine acid began. The coal operator has had to deal with the acid water produced in his mines, and his losses from corrosion of pumps and equipment have been tremendous. Other industries and the public have been forced to spend large amounts for their water requirements because of the acid discharge from the mines. There are many contradictions in the literature on this subject probably because of variations in the analytical procedures followed and differences in the methods of calculation and reporting. The opinions previously expressed may be summarized as follows: The production of mine acid has been looked upon as a simple chemical reaction wherein the oxygen of the air has reacted with pyrite to produce ferrous sulphate and free sulphuric acid. This reaction appears in many chemistry texts and generally has not been considered worthy of further study. Since the measurement of hydrogen ion concentration or activity was developed and expressed in terms of pH, indicator and instrumental methods for pH determination have become readily available, making it natural to think of it as a measure of acid content. Special purpose analytical methods applicable to potable waters, sewage effluents, and boiler waters, based in general upon the determination of what may be regarded as impurities, have been applied to more concentrated solutions constituting mine discharges, without reference to the fact that these specialized procedures had been developed to serve as operating guides for water-treatment plants. It has been assumed that the pounds of acid deliv- Table I. Showing pH and Titratable Acid of Varying Concentrations of Ferrous and Ferric Sulphates Ferric Ferrous ion Acldfty Ion Acidity (Fe+++) ppm CaCO3 We++) ppm CaCOa ppm PA equiv. PPm pH equiv. 11.200 1.80 30,000 11,200 3.98 20,000 5,600 2.01 15,000 5,600 4.15 10,000 2,800 2.21 7,500 2,800 4.41 5,000 1.120 2.48 3,000 1,400 4.68 2.500 560 2.60 1,500 140 5.00 250 280 2.65 750 28 5.00 50 112 2.81 300 56 3.07 150 28 3.38 75 ered by a mine was roughly a constant varying with the number of acres mined, and that the concentration of the acid in the drainage was inversely proportional to the volume of flow. Increased flow was thought to result in a dilution of the acid produced. When the Sanitary Water Board of the Commonwealth of Pennsylvania established a fellowship at Mellon Institute, the first approach to the problem was to review some of the chemical reactions involved and their relation to acid production. Since the sulphates of iron were among the principal constituents of coal-mine drainage, the first studies were based upon solutions of chemically pure ferrous and ferric sulphate. These are both neutral salts, that is, the iron and the sulphate are equivalent, but they produce acid reactions in water solution because as salts of strong acids with weak bases, they hydrolyze to form acid solutions. Solutions of each salt, of varying concentrations, were prepared and the pH, titrable acidity and iron and sulphate

Jan 1, 1952

By E. E. Anderson, R. L. Newacheck, J. Kohl

Radioactive materials were used underground before the advent of the nuclear reactor. Radium containing ores have been mixed with cements and other materials for later location in oil wells by gamma ray logging since 1940.1 Perforating bullets have been marked with radium salts,, and radioactive materials have been added to drilling muds to find the cause of "lost circulation."&apos; More recently colloidal suspensions of radioactive cobalt-60 and silver-1103 and solutions of 1-1314 have been used in waterflooding operatiorns to determine injection profiles, flow paths and water velocities. In this paper a recent and interesting underground tracer application is discussed. Development work and field engineering were carried out under the direction of Tracerlab, lnc., with assistance from personnel of the client. A gaseous tracer, argon-41, was used in an attempt to determine the cause of leakage from an underground gas storage reservoir. STATEMENT OF PROBLEM This gaseous tracer application was carried out for the Natural Gas Storage Co. of Illinois. This company has developed an underground gas storage field 1,750 ft below the surface of the earth near the village of Herscher, Ill., about 60 miles south of Chicago. The total volume of gas stored in the field was estimated to be about 13 billion cu ft in the spring of 1954. The field has a maximum estimated storage capacity of 90 billion cu ft. The gas is contained in the Galesville sandstone, previously a water bearing formation of about 25 per cent porosity. The water as it is pushed out by the gas forms a barrier for the gas around the edge of the reservoir. Over the top and beneath the bottom of the reservoir are located several layers of impervious rock, which barring cracks or unplugged and unknown wells should act as seals. Twenty-one injection and withdrawal wells were drilled into the Herscher reservoir and gas injection was begun during the summer of 1953. After a few weeks of gas injection, evidence of gas leakage was detected at the surface; water wells began bubbling and losing prime, and dry spots in the fields indicated gas leakage from uncharted oil wells. In every case, the leakage could not be traced back any further than the water bearing formations a few hundred feet below the surface. An extensive search was made to locate uncharted oil wells which were believed to be possible channels for the gas from the storage formation to the water containing formation. Long-time residents in the area assisted in the location of several abandoned oil wells. The wells when located were cleaned out and either used as vent wells to relieve the gas pressure in the water formations, or were plugged with cement. None of the wells even approached depths which would endanger the soundness of the storage reservoir. All of the injection wells were tested for leaks in the casings by conducting neutron surveys and by other techniques. To detect any leakage up through the cement-filled annulus surrounding the casing, a section of the casing wall in one well was milled out. Then, plugs were installed above and below the cut-out section and pressure applied to the annulus, which proved conclusively that the cement was sound in that well. The well was then plugged with cement and abandoned. In another well, ammonia was injected into the well and a temperature log was made. It was hoped that a leak would be re-

Jan 1, 1956

By M. A. Selim, W. H. Somerton

Thermal expansions and heats of reaction of three typical sandstones were measured in the temperature range of 25°to 1,000°C. The significance of these data in subsurface heat-transfer calculations is discussed. Linear thermal expansions were measured both parallel and perpendicular to the bedding planes. Volume expansions are reported as the perpendicular expansion plus two times the parallel expansion. Expansion behavior of the sandstones was found to be controlled by the expansion characteristics of the quartz content. Differential expansion of the quartz grains and other minerals included in the sandstones caused permanent deformation of the heated samples after they were cooled to room temperature. Structural damage resulting from heating is probably an important cause of the reduction of thermal conductivity of heated samples. Measurements of heats of reaction were based on the known heat required for a- quartz inversion. Thermal reactions, which probably include deby-droxylation of clay minerals and decomposition of carbonate minerals contained within the samples, were found to require more than one-fourth of the amount of heat necessary to raise the temperature of the rock alone. In shales and limestones, the reaction heat could be substantially greater than that required from specific heat considerations alone. INTRODUCTION In earlier work, l-3 methods of measuring thermal properties were developed and thermal data for several sedimentary-rock types were reported. Data included thermal conductivity, specific heat and thermal diffusivity. In addition, calculations of heat transfer in subsurface formations may require data on the thermal-expansion characteristics of rock and the heats of reaction of mineral constituents. Thermal expansion of rocks is relatively small in magnitucle and, from the standpoint of change in bulk density, has only minor effects on heat-transfer cliaracteristics. Thermal-expansion behavior, however, may have significant effects on the structure of rocks. Differential expansion of different minerals and along different crystal-lographic axes may result in structural damage which could effectively reduce the thermal conductivity of the rock. Many mineral constituents of sedimentary rocks undergo phase or related changes when heated to sufficiently high temperatures. Although some of these reactions (such as the a-ß quartz inversion) are reversible and the heat absorbed is returned to the system upon cooling, many of the major reactions are irreversible. Through certain temperature ranges, the additional heat energy required to complete the reactions may be nearly as great as that necessary to raise the temperature of the rock in the absence of thermal reactions. If the heat requirements for these reactions are not considered, serious errors in heat-transfer calculations may result. Thermal-tzxpansion and thermal-reaction data have been obtained for three typical sandstones. Methods of measurement and results obtained are reported and discussed in the following. THERMAL EXPANSION OF SANDSTONES The linear expansions of three outcrop sandstones (Bandera, Berea and Boise) have been measured in the temptbratue range of 25° to 1,000° C. Expansion measurements were made in the directions parallel and perpendicular to the bedding. The differential expansion apparatus used is that described by Mitoff and Pask.4 The test sample is heated in an electric furnace with a temperature rise of 6°C/minute. The lengthening of the samples upon heating is compared with the small and known expansion of a fused silica rod. The change in length is transmitted to an X-Y recorder by means of a Statham transducer. A maximum error of ± 1.5 per cent has been obtained with this apparatus

By Robert M. Asimow, Bisuddhi Riddhagni

THE use of back-reflection Laue photographs for crystal orientation is standard in many research projects. The time required for the orientation of any given crystal depends largely on luck and the skill of the investigator. In any case, where large numbers of crystals are used, this time can be quite considerable when the standard methods involving a Greninger chart and Wulff net are employed. For this reason the technique described below was developed. This new procedure applicable for cubic crystals (without preparation of new charts) largely eliminates the skill and luck factors and makes possible the determination of orientation in minutes. The procedure works in the following way: the Laue photograph is examined and arbitrarily an important hyperbola is picked out. An important hyperbola is defined as one which contains some spots through which several hyperbolas pass. Such a hyperbola inevitably corresponds to a low-index zone axis. The angles between the reflection planes causing all the spots on this hyperbola can be read directly from a Greninger chart in the usual manner. If the hyperbola contains more than five spots, these interplanar angles will positively identify the zone axis and the indices of each of the reflecting planes. However, initially since neither the zone axis nor the Miller indices of a diffraction spot are known, there are a very large number of possibilities that must be considered. Using tables of interplanar angles to check out every possibility would be quite tedious, so that this approach is not generally used. Instead, various poles are plotted on a stereographic projection. In the cubic system, however, it is possible to plot out once and for all the angles between all planes of a particular zone axis up to some upper limit of Miller indices. This has been done for all low-index zone axes in Fig. 1. Given these charts, it is quite a simple matter to identify the spots of the unknown hyperbola. All that is necessary is to plot the angles read off the Greninger chart to the same scale as Fig. l(b) on a strip of paper. Then this strip can be slid along each zone axis of Fig. l(a) until a good fit is obtained. Once this is accomplished, the Miller indices of each spot are given and thus the orientation problem is solved. To check, it is possible to pick a second hyperbola and see whether a compatible solution is obtained. This is most conveniently done by picking for the second hyperbola one that contains a spot common with the first. If the specimen axis is chosen parallel to the incident X-ray beam, then usually the [ loo] hyperbola will appear on the film as one of the more important hyperbolas and consequently will be either chosen initially or as a check. Using the [100] hyperbola, after the (001) or (011) spots have been identified, the latitude and longitude of the specimen axis can be read directly from the Greninger chart so that it is unnecessary to use a Wulff net for the entire orientation procedure. This makes the operation not only quicker but more accurate. Of course, if neither the first nor second hyperbolas correspond to [loo] then a third hyperbola can be chosen.

Jan 1, 1967

By J. C. Rae

A Japanese mining engineer was given a year&apos;s leave of absence by his company and dispatched on a tour of inspection of all the principal coal mining regions of the world. He visited the New Castle region of England, Wales, Scotland, France, Germany and Alsatian mines, the underseas col¬lieries at Nova Scotia and Vancouver, and our own important fields of Illinois, Pennsylvania, West Virginia, Kentucky, Colorado, Utah and Wyoming. You are probably surprised to know that after seeing all these, he pronounced the Gebo mine of the Owl Creek Coal Company as the most interesting operation he had seen. The Gebo Mine is located about three miles west of Kirby on the Chicago, Burlington & Quincy Railroad in the Big Horn Basin region of Wyoming, 500 miles north of Denver and 200 miles south of Billings, Montana. The mine is working on a coal bed five or six to eleven or twelve feet&apos; thick, dipping 22 to 23 degrees toward the northeast. The coal occurs in what is called the Eagle formation; and is found between two massive sandstones from 25 to 65 feet thick, which, weathering into bold&apos;, gray-colored cliffs, serves as a guide to the location of the coal bed. The deepest workings have an overburden of 800 feet. Cleavage planes of the coal extend at an angle of 45 degres to the pitch, and perpendicular to roof and floor. The coal itself is generally clean and not difficult to cut. It ignites readily from spontaneous combustion, however, which circumstance makes an ever-present fire hazard which must be carefully watched and guarded against. The coal is classed as sub-bituminous with analysis about as follows: Moisture 17% Volatile Matter 33% Fixed Carbon 46% Ash 4% Sulphur, about ½ % B. t. u. value a trifle under 11,000. It is known as Owl Creek coal, is one of the best domestic and steam fuels in the West, and is very popular in Montana, Idaho, Washington, Wyoming, South Dakota and Nebraska. While the coal slacks somewhat rapidly when exposed to the weather, it stores readily for months if placed under cover. On account of the somewhat isolated position of the mine as regards manufacturing establishments, there was a decided difficulty in disposing of slack, especially during the early years of the mine (1907 to 1914), and approximately 10 per cent of the entire product had to be wasted as slack during those years. This difficulty has been largely overcome since 1914, due partly to growth of manufacturing establishments in tributary territory, and also to enlarged transportation facilities. Also largely due to its isolated position, the mine has been compelled to use an extremely large proportion of box cars, much over 50 per cent of the product (in some years 75 per cent) being shipped in box cars, which, of course, entails extensive loading equipment and increases cost of loading. The output of the Gebo Mine for 1920 was 348,000 tons, for 1921, 273,000 tons; the rate of production during the year 1922 was over 1,700 tons daily, and 2,000 tons daily in January, 1923. The mine was developed by driving three openings straight down the pitch, the main slope being 12 to 14 feet wide and 3,000 feet long. A little distance from the surface a pair of level entries were turned right and left, known as the first East and first West. From these, rooms were driven up the pitch and chute mining attempted with only indifferent success: The failure of this rise-room method of mining was attributable to several

Jan 1, 1923