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By J. E. Campbell, G. D. Calkins, N. M. Ewbank, M. Pobereskin, A. Wesner

GRINDING for size reduction affects the economics of many processes and products. It is essential as the first step in many industrial processes and is also a finishing step for materials with properties depending on particle size, such as talc, cement, and silica sand. Intermediate and fine grinding are vital operations in the U. S. cement industry, which is producing more than 250 million bbl of cement per year.' Wear of the grinding media is a large part of the grinding operation cost. Problems encountered in grinding cement are so complex that evaluation of efficiency and economy of grinding media is difficult.2 It has been especially difficult to evaluate the relative effectiveness of different types of balls because there are no good testing techniques. Many other industrial operations can be evaluated on a laboratory scale with reasonable accuracy. This does not hold true for evaluation of grinding balls. The consistent results obtained in a laboratory test under a given set of conditions are not always borne out in field application. Rough evaluations of the effectiveness of various compositions and types of grinding balls have been made in the field by using a full charge of one type in a mill and comparing the production record with another run using another type of ball. This method is time-consuming and not very precise, as the second run may not have been carried out under identical conditions. Laboratory-scale tests, on the other hand, have yielded inconclusive results, and many investigators have turned their attention to the development of a field testing technique. Field testing small sample lots of grinding balls has been impractical because it is difficult to identify and recover the test specimens from the grinding mill, and individual groups of balls that have undergone different heat treatments can not be separated.".4 To overcome these difficulties, previous investigators have identified the balls by distinctive marks, notches, and drilled holes, but this procedure has three serious drawbacks: 1) Grinding characteristics and quality of the steel balls may be affected. 2) Physical markings may be worn away in the grinding process, especially during a prolonged run. 3) Recovery from the bulk of the charge will be extremely difficult because the markings are hard to see and may be masked by a coating of the product. To circumvent these difficulties, a radioactive-tracer technique was proposed for recovery and separation of steel grinding balls and subsequent evaluation of the various compositions of the balls. The proposed technique involved five basic operations: 1) Thermal-neutron irradiation activation5 of each group of test grinding balls to a different level of specific radioactivity. 2) Addition of groups of radioactive steel-ball specimens into a ball tube mill. 3) Recovery of radioactive steel-ball specimens from the bulk of the mill charge. 4) Separation of the various groups by their specific radioactivity. 5) Evaluation of actual grinding ball wear. Before any physical tests were performed, required neutron irradiation intensity and time were calculated. Probable composition of the steels to be used was ascertained. An examination was made of the radioactive nuclides8 to be formed which would contribute measurably to the radiation level immediately after irradiation and during the test operation. The radioisotopes formed, their types of radiation, and their half lives are listed in Table I. Of these radioisotopes only iron-59 and chromium-51 were significant for the actual wear test. The intensity of radiation that could be detected by a Geiger counter when the test was completed was the basis for the minimum activation level established. The intensity of radiaton that could be safely handled at the beginning of the test was the basis for the maximum activation level, although this was not considered a major problem. Ten groups of grinding balls of various composition and/or surface or heat treatment were to be tested. One group was designated for the minimum irradiation time. The remaining groups were designated for irradiation periods that increased by increments of 33 pct from that of each preceding group. This difference was considered enough for separation and identification of the groups by comparison of specific activity. Potential Hazards: Possible radiation hazards that might be encountered during this experiment were evaluated for the three important phases: 1) the radiation hazard of placing balls and removing them from the mill, 2) contamination of the product cement by radioactive material worn from the balls, and 3) contamination of the steel by the radioactive balls left in the mill. The radiation intensity expected from the whole group of radioactive balls was calculated to be 250 milliroentgen per hr at 1 ft. This meant the balls would require special shielded packaging and warning labels on the shipping containers. In a radiation field of 250 mr per hr a man can work for 1 hr without exceeding maximum permissible weekly exposure. Since the balls could be dumped into the mill in a matter of seconds, relatively little radiation exposure was anticipated at this stage of the operation. If the weight loss in the balls was 7.7 pct per month and the cement feed through the mill was

Jan 1, 1958

By Kerstin Malmqvist

A prospecting project for deep-seated sulphide ore bodies will be expensive and means a considerable economic risk. The high costs depend on the demand for extensive drilling. It is therefore important to find effective methods to estimate the outcome of different prospecting strategies. The results from a preliminary study of the possibility of using mathematical methods for this purpose are presented. We have found that mathematical methods, especially simulation technique, might be an important tool in such exploration work. A mathematical model of an ore field and of a drilling procedure has been constructed. A set of parameters describes the geometry, the size distribution and the intensity of ore bodies. Another set describes the geometry and the intensity of drill holes. Some preliminary results are given of Monte-Carlo simulations and probability calculations.

Jan 1, 1977

By M. L. White, J. L. Shafer, C. L. Caenepeel

Often an in situ leach is the only practical economic method for copper recovery from small low grade oxide deposits. The decision to develop a copper property by an in situ blast and leach is strongly dependent on ore grade, tonnage, hydrology, and the copper extraction and acid consumption rates. The inherent leachability of the ore will in part determine the desired amount of fracturing of the ore deposit by the blast in order to obtain a copper extraction rate that is economical. For the last several years Occidental Research Corp. (ORC), has undertaken a modest effort to address the question of how to best identify and determine the information necessary to predict and control the operation of an in situ leach for the economic recovery of copper.

Jan 2, 1979

By I. L. Nichols, D. C. Seidel, D. E. Traut

A fluidized countercurrent ion-exchange system was developed, operated, and evaluated by the Bureau of Mines. The system consisted of integrated multiple-compartment absorption and elution columns in which the solution flows are continuous except for short periods when resin increments are withdrawn. The exchange of uranyl sulfate between a simulated acid leach uranium solution and a strong-base, ion-exchange resin together with the subsequent elution with an acidified sodium chloride solution was studied. The effects of the number of sections, section heights, amount of resin withdrawal, and solution flow rates were investigated. Also examined were the kinetic and equilibrium relationships for the absorption step. The experimental data indicate a strong interdependence between variables. Solution retention time appears to be a major limiting variable. The column was efficient over a range of conditions, but close control was needed for optimum operation.

Jan 1, 1977

By J. L. Bruce, G. W. Nicolson

THE country rock of the Mavrovouni mine of the Cyprus Mines Corp. is hydrothermally altered, disintegrated pillow lava, with very little tensile strength ("short" ground). In places, especially when wet, it is heavy ground, though not extremely heavy except where movement has been caused by underlying or nearby stoping operations. Damage by caving from roof or sides of galleries is not sudden but progressively faster when the size of the cave increases as a result of the fall of small pieces. CEMENT AND CONCRETE LININGS During early operations, galleries in the country rock were supported by customary mine-timbering methods. Available timber, however, was of poor quality and of relatively high price, and replacements were frequent and expensive. Furthermore, the massive sulphide ore bodies required large volumes of fresh air for ventilation and cooling, and it was therefore desirable to provide smooth-lined galleries with low resistance for the introduction of air to the ore bodies. It was also desirable to eliminate hazards of fire in shafts and underground workings insofar as this would be practical. Cement and concrete were relatively cheap, and plans for using these in support of shafts, underground galleries, raises, and chutes, were developed steadily. Circular shafts with diameters ranging from 10 to 15 ft were adopted for all permanent shafts and airways. In places these were lined with concrete poured behind sectional pressed-steel forms especially constructed for this purpose. In other places, they were lined with precast-concrete blocks. It was found that such precast concrete of good quality could be produced on mass production scale and delivered at the mine from most favorable point of manufacture, at a cost of about 25 cents per cubic foot of concrete. This refers to prewar costs. As the use of concrete blocks in circular lined galleries was developed, it was found that there were very great advantages in the reduction of air resistance as compared with customary timber supports. After best methods had been developed, it was found that breakage and replacements were extraordinarily low, averaging only about 1 ½ pct per year during a period of about 15 years. Furthermore, it has been possible to reclaim for re-use a very large percentage of the concrete blocks from abandoned galleries. During the history of the operations precast-concrete blocks have been used for lining shafts, raises, ore chutes, galleries, and shaft stations, ranging from 3 to 16 ft id. The thickness of lining has varied from 5 to 10 in.; with I2 in. thickness in a few cases. Junctions of two galleries at angles ranging from about 30 to 60° are usually made of poured concrete behind sectional forms, or of a combination of poured concrete and precast blocks. In nearly all cases it has been found advisable to make the point of the wedge at the junction, of

Jan 1, 1947

By Paul D. Merica

One of the most useful, and at the same time least commonly used, methods of thermal analysis for the determination of transformations in metals and alloys consists in the recording of the time intervals required for successive increments of temperature change during heating or cooling, the temperature of the furnace containing the specimen being altered at a uniform rate. The curve obtained by plotting these time intervals as a function of the mean temperature of the specimen durnig the interval is called the inverse-rate curve. It is probably due to the fact that no simple and convenient method has apparently been available for the measurement of the successive time intervals that this method has not been so generally used as, for example, the differential method, for which several types of automatic or semi-automatic apparatus have been designed. Whenever this method has been used, the intervals have usually been measured with the use of a chronograph; its operation as practised at the Bureau of Standards1 is as follows: The temperature of the specimen is measured by a thermocouple and a dial potentiometer. The operator sets the potentiometer at successive values of the electromotive force differing by equal increments, usually 0.02 millivolt, and records the exact instant on the chronograph, by pressing a contact key, at which the galvanometer coil passes through its null position. Two-second intervals are also recorded on the chronograph record, and the number of seconds elapsing between successive signals is afterward counted from the record and plotted as a function of the electromotive force or of the temperature. This is an admirable method and by it most minute thermal arrests may be detected, but it requires a good chronograph, which is generally difficult to obtain, and the time of one operator during the recording and, subsequently, to read the intervals from the record; the latter operation often takes from one to two hours. There is here suggested a simple, convenient, and, it is believed, equally accurate method of recording the successive time intervals, by which the

Jan 1, 1921

By N. De Nevers

A calculation method has been developed for carbonated water flooding. This method takes into account the effects of oil viscosity reduction and oil swelling, due to carbon dioxide transferred to the reservoir oil from the carbonated flood water. It shows the effect of changes in carbonation pressure and carbonated water slug size. The method is based on a Buckley-Leverett-type linear flow model. The mathematical approach is similar to that developed by Welge, et al, for enriched gas drives. However, it contains a very important advance over the previous papers in its solution to the problem of injection of a "slug" of carbonated water rather than continuous injection of fluid of constant composition. The method also shows the chromatographic nature of the transport of Co 2 in a carbonated water flood. Sample calculations indicate that viscosity reduction is the most important source of incremental oil recovery. Swelling contributes to a lesser extent. Raising the carbonation pressure increases recovery, but this effect diminishes as pressure is increased. increasing the slug size also increases recovery, but again this effect diminishes as slug size is increased. INTRODUCTION In recent years there has been considerable interest in carbonated water flooding as a secondary recovery method. Several carbonated waterflood field tests are now in progress.1'3 Yet there is no published method for predicting the results of a carbonated water flood. This paper presents a method for predicting the oil recovery for ; carbonated water flood. In a carbonated water flood. a carbonated water slug is injected into the reservoir. Then this slug is driven forward by injecting plain water behind it. The field operator must choose the carbonated water slug size and the CO2:water ratio in the slug. The calculation method presented here shows the effects of both slug size and CO2:water ratio on oil recovery. The calculation assumes that the addition of carbon dioxide to flood water increases the oil recovery by two mechanisms: 1. Carbon dioxide transfers from the flood water to the oil, thereby reducing the viscosity of the oil. At the lower viscosity, the oil flows more readily and is more easily swept out by flood water. 2. The carbon dioxide which transfers to the oil phase causes the oil phase to swell. Assume that the residual oil saturation (volume per cent of pore space) is the same whether the oil is carbonated or not. Then the same volume will contain less stock tank oil if it is carbonated, due to oil swelling. The plain water driving the carbonated water slug forward will strip the CO2 out of this residual oil., causing this oil to shrink. This swelling-followed-by-shrinking leaves behind a lower residual oil saturation than would have been left behind by a plain water flood. Since some oils swell to almost twice their original volume at reasonable CO2 pressures, this effect may be large. Two additional benefits sometimes claimed for carbonated water flooding are: 1. Adding CO2 to flood water greatly increases the water injectivity in some cases.1 (Injectivity = BPD of water injected/ft of sand x differential bottom-hole pressure, psi.) CO2 forms a weak acid in water; injecting carbonated water is, therefore, somewhat like acid treating. If a formation responds to acid treating, it will probably have a higher injectivity for carbonated water than for plain water. Although the increased injectivity increases the profitability of a carbonated water flood over a plain water flood, it does not change the production curve based on total fluid injected. Therefore, the method presented here does not show any effect of increased injectivity. 2. It is claimed in one patent4 that carbon dioxide will react with some constituents of the oil to form detergents. These detergents are expected to increase the oil recovery. There is little data available to support this idea. The calculation method presented here takes no credit for extra oil recovery due to the formation of any such detergents.

Jan 1, 1965

By Neil H. Cole

Cost and price inflation and financial gearing through loans are characteristic of modern resources projects. Conventional discounted cash flow and 1P.R analyses, in real terms, and without financing, have inadequacies which can yield misleading results. The degree of improvement in value for project equity owners is limited to the value of the additional tax benefit accorded through the financing instrument used. Debt financing may be low in price and value, but its priority means that the equity owners assume a higher risk, thereby increasing the cost, or price, of their capital. The cost of equity capital, or investment hurdle rate, is related to the Capital Asset Pricing Model. Distortions do occur, with inflation, due to depreciation effects, which can cause the adoption of misleading project rate of return or hurdle rate criteria. For investment decisions, the greatest accuracy possible in forecasting is essential, especially for inflation rates,

Jan 1, 1985

By J. W. Leonard

One of the more obvious phenomena which relate to coke formation is that soft coals of low volatile matter content tend to yield hard or high strength cokes while hard coals of high volatile matter content tend to yield soft or low strength cokes. This observation led researchers of the Coal Research Bureau of West Virginia Univ. to study the possibility of using Hardgrove grindability as a supplemental rank indicator for coking properties. Grindability has been used for many years as a means to measure the resistance of coal to comminution but it has rarely if ever been considered for evaluating coals for carbonization. The coal proper- ties which are measured by grindability are complex, but recent work with the difficult-to-correlate high volatile coals has demonstrated that a 93% correlation exists between grindability and the total petrographic rnicrinites, exinites, and resinites. This work demonstrated that inherent ash had no apparent effect on the grindability of high volatile coals, suggesting that the coal substance is apparently harder than the associated ash. These observations, together with the patented Burstlein process," indicate that the distribution of hard and soft components in the individual high volatile coals and blends of coals may be quantitatively interrelated to coke stability through selected ranges of coal volatile matter. The objective of this progress report is 1) to demonstrate the statistical relationships which exist between coal grindability and the numerous coal petrographic entities and 2) to present the results of initial studies concerning the correlation of the coke physical strength index of stability with coal volatile matter and grindability.

Jan 3, 1964

By M. E. Fine

IN the apparatus for measuring internal friction shown in fig. 1, a modification of that described by Kê,1,2 the parasitic energy loss or background is equivalent to a Q-1 of approximately 4x10-5. This low value of background was obtained by using a jeweler's collet of appropriate size for the upper support, eliminating lateral motion without immersing the bob in oil, and designing the bob to reduce air resistance. The pendulum is supported on the frame through a snug fitting, sliding rotary joint (fig. I-B). A slight twist of the top imparts essentially pure torsional motion to the pendulum. No evidence of energy loss in the sliding joint has been observed. In the bob (E) steel balls are used in the torsion arm, and the bottom is filled with lead giving a total weight of 94 g. A lighter bob may be used for small diameter specimens. Another bob was a solid disc with provision for adding discs to change the frequency. The oscillating image of the filament is recorded on a revolving drum camera (fig. 1-3) driven by a synchronous motor. This allows use of frequencies higher than can be measured visually on a glass scale. Samples varying in diameter from 0.020 to 0.075 in. have been measured, the frequencies have ranged from 1 to 12 cps, and the maximum strain varied from 0.5 to 10x10-5. With appropriate modification the sample may be placed in a thermostat. The apparatus may also be used to measure the shear modulus. In the examples of measurement given in table I the internal friction of Permendur and the 46.8 Ni-Fe alloy are due, for the most part, to magneto-mechanical hysteresis8 (elastic hysteresis accompanying stress-induced magnetostriction) which depends on strain amplitude and increases with annealing. The internal friction of most materials decreases with annealing. References IT. S. K6: Physics Review (1947) 71, 533. 2 T. S. Kê: Transactions AIME (1948) 176, 448; Metals Technology (June 1948) TP 23703. 3 R. Becker and W. Döring: Ferromagnetismus. 365 (1939). J. Springer.

Jan 1, 1951

By F. W. Sperr

The byproduct coke oven is the most important artificial source of fuels for metals melting. Its products are solid, liquid, and gaseous in form. The amount of coke and primary byproducts obtained per ton of coal varies with the nature of the coal. The following represents the results obtained in many plants: Metallurgical coke (over 3/4 in.).............................. 69 per cent. of coal Domestic coke (1/2 to 3/4 in.)................................ 2 per cent. of coal Breeze (under 1/2 in.)...................................... 4 per cent. of coal Surplus gas, cubic feet per net ton........................... 6600 Gas for heating ovens, cubic feet per net tor i................. 4400 Tar, gallons per net ton.................................... 9 Ammonium sulfate, pounds per net ton....................... 25 Benzols, gallolls per net ton................................. 4 All of these products, except ammonium sulfate, have important fuel value. Part or all of the metallurgical coke may be crushed or screened to furnish additional domestic coke. If producer gas is used, there will be a surplus of 11,000 cu. ft. of coke-oven gas per net ton of coal. The use of coke in the metallurgical industries has been retarded by the mental attitude of writers who disregard the dependence of coke quality on the conditions of its manufacture. Terms like "byproduct coke," ('beehive coke," "48-hour coke," and "72-hour coke" are only generally descriptive and tell nothing of the coke quality. For example, one writeri stated that in brass-foundry practice, Connellsville coke is considered rnuch superior to byproduct coke on account of the high sulfur content of the latter. Another2 says, "At equal prices for hard Connellsville coke and for byproduct coke at the foundry door, the hard coke is unquestionably cheaper. It will melt more metal per ton than the byproduct coke and requires less labor, as the byproduct coke is less dense and burns faster, thus requiring more frequent firing." A little study of the conditions under which coke is manufactured will show the error in making general statements like these. The percentage of sulfur in coke depends largely on that in the original coal, and byproduct coke from a

Jan 1, 1921

By R. L. Manegold

This paper outlines the reasons for and the application of suspended magnets above belt conveyors for the purpose of removing tramp iron to protect equipment. The design of a magnet strong enough to extract iron through a distance of 24 in. is described along with the proper method of application in the field. Advantages of the suspended magnet in contrast to the conventional magnetic pulley are also discussed. IN RECENT years there has been a decided trend toward bringing ores and coal out of open-pit and underground mines by long, sloping single-stage belt conveyors. Because the high investment cost of haulage conveyors dictates maximum speed and full loading for economy, conveyor belts are running faster than ever before, with a greater burden on the belts. In addition, larger blocks of ore are being mined and conveyed. Rougher use of large mining equipment generates more tramp iron; mechanical loading rejects nothing. Therefore the extraction of tramp iron and other deleterious materials for the protection of haulage conveyors at transfer points, chutes, and idlers is becoming an increasingly difficult job. In our experience, the best method of protecting long haulage conveyor belts is to extract the tramp iron, at least the large tramp iron, ahead of the main haulage conveyor or series of conveyors. It is most practical to collect the tramp iron at the earliest source possible, usually directly after the primary car unloading dump or after the primary crushers, which are ordinarily not affected by tramp iron, see Fig. 1. The coal or ore from its primary source is fed by chutes or feeders onto a buffer or shock belt conveyor, a short, ruggedly built conveyor system that can be considered somewhat expendable. Its primary function is to bring the ore up to speed for transmission to the haulage conveyor. There is no assured method of protecting this conveyor from tramp iron damage at its feed end. Likewise, there is a possibility, although remote if the magnet applied to this belt is installed correctly, that the tramp iron extraction process will damage the shock belt, which is, however, short and comparatively inexpensive. Magnetic head pulleys with their maximum effective range of about 8 or 10 in. cannot be considered for large installations, even though they are desirable because of the continuous and automatic tramp iron removal feature. Further, the most dangerous form of tramp iron with respect to potential damage to a haulage conveyor is the long rod, roof bolt, or rail segment, which is not extracted satisfactorily because of the limited tangential contact of such an iron shape with the magnet. Magnetic head pulleys are also undesirable because the dribble or fines from the head pulley discharge at the same point as the tramp iron. Screens must be used to separate the fines from the magnetics, resulting in additional expense and loss of head room. Magnetic detectors have been used with success, but they only indicate the presence of iron, and the conveyor belt must be stopped for its manual removal. The stoppage of any conveyor means that all preceding equipment must be stopped or a cascade of material on the idle conveyor will result. Occasionally combinations of detectors and magnetic head pulleys are used. The detector precedes the magnetic head pulley and detects only large pieces of tramp iron, while the magnetic pulley subsequently takes out the smaller ferrous metals. Suspended magnets overcome the disadvantages of magnetic head pulleys, detectors or a combination of both. There are several magnetic extraction problems involved in applying a suspended magnet at this

Jan 1, 1953

By R. L. Manegold

This paper outlines the reasons for and the application of suspended magnets above belt conveyors for the purpose of removing tramp iron to protect equipment. The design of a magnet strong enough to extract iron through a distance of 24 in. is described along with the proper method of application in the field. Advantages of the suspended magnet in contrast to the conventional magnetic pulley are also discussed. IN RECENT years there has been a decided trend toward bringing ores and coal out of open-pit and underground mines by long, sloping single-stage belt conveyors. Because the high investment cost of haulage conveyors dictates maximum speed and full loading for economy, conveyor belts are running faster than ever before, with a greater burden on the belts. In addition, larger blocks of ore are being mined and conveyed. Rougher use of large mining equipment generates more tramp iron; mechanical loading rejects nothing. Therefore the extraction of tramp iron and other deleterious materials for the protection of haulage conveyors at transfer points, chutes, and idlers is becoming an increasingly difficult job. In our experience, the best method of protecting long haulage conveyor belts is to extract the tramp iron, at least the large tramp iron, ahead of the main haulage conveyor or series of conveyors. It is most practical to collect the tramp iron at the earliest source possible, usually directly after the primary car unloading dump or after the primary crushers, which are ordinarily not affected by tramp iron, see Fig. 1. The coal or ore from its primary source is fed by chutes or feeders onto a buffer or shock belt conveyor, a short, ruggedly built conveyor system that can be considered somewhat expendable. Its primary function is to bring the ore up to speed for transmission to the haulage conveyor. There is no assured method of protecting this conveyor from tramp iron damage at its feed end. Likewise, there is a possibility, although remote if the magnet applied to this belt is installed correctly, that the tramp iron extraction process will damage the shock belt, which is, however, short and comparatively inexpensive. Magnetic head pulleys with their maximum effective range of about 8 or 10 in. cannot be considered for large installations, even though they are desirable because of the continuous and automatic tramp iron removal feature. Further, the most dangerous form of tramp iron with respect to potential damage to a haulage conveyor is the long rod, roof bolt, or rail segment, which is not extracted satisfactorily because of the limited tangential contact of such an iron shape with the magnet. Magnetic head pulleys are also undesirable because the dribble or fines from the head pulley discharge at the same point as the tramp iron. Screens must be used to separate the fines from the magnetics, resulting in additional expense and loss of head room. Magnetic detectors have been used with success, but they only indicate the presence of iron, and the conveyor belt must be stopped for its manual removal. The stoppage of any conveyor means that all preceding equipment must be stopped or a cascade of material on the idle conveyor will result. Occasionally combinations of detectors and magnetic head pulleys are used. The detector precedes the magnetic head pulley and detects only large pieces of tramp iron, while the magnetic pulley subsequently takes out the smaller ferrous metals. Suspended magnets overcome the disadvantages of magnetic head pulleys, detectors or a combination of both. There are several magnetic extraction problems involved in applying a suspended magnet at this

Jan 1, 1953

By J. S. Crump, C. R. Hocott, A. E. Hoffman

Planning of the efficient operation of a gas-condensate reservoir requires a knowledge not only of the gross phase behavior of the system but also of the equilibrium distribution of the various components between the gas and condensate phases. This equilibrium distribution can be calculated with appropriate equilibrium constants. In this paper are presented equilibrium constants determined experimentally for the oil and gas phases initially present in the same reservoir and for the gas and condensate phases of the gas cap material at a series of pressures below the original reservoir pressure. Also presented is a method for the correlation of the experimentally determined equilibrium constants. The utility of the correlation is demonstrated further by an example of its use in adjusting the equilibrium data to permit their application to another gas-condensate system of similar composition. INTRODUCTION Planning of the efficient operation of a gas-condensate reservoir requires a thorouugh knowledge not only of the gross phase behavior of the particular hydrocarbon system but also of the equilibrium distribution of the various components between the gas and condensate phases. At the initial conditions of reservoir temperature and pressure, the original hydrocarbon materials in a gas-condensate reservoir or in the gas-cap of an associated reservoir exist in a single, homogeneous vapor phase. However, some condensation of hydrocarbons to a liquid phase usually occurs in the reservoir as pressure declines incident to production. Because of this condensation. the produced gas changes composition continuously. The composition of the produced gas, as well as that of the condensed liquid, can be calculated from the composition of the original reservoir material through the application of appropriate equilibrium constants or "K"-values provided these values are known. Equilibrium constants have been used for many years in problems of surface separation of gas and oil and natural gasoline recovery. However, no satisfactory equilibrium data at reservoir conditions have been available for the higher boiling hydrocarbons which acquire abnormal volatility at high pressure and are therefore present in the gas phase in condensate reservoirs. Early work in this field indicated that the behavior of the higher boiling hydrocarbons largely determines the behavior of gas-condensate systems at high pressure.' Consequently, a portable test unit was designed to permit determination of the distribution of the heavier hydrocarbons between the gas and liquid phases of a gas-condensate system. This paper describes the test equipment and presents the results of an investigation of the reservoir fluids from an associated oil and gas reservoir in the Frio formation in Southwest Texas. Also presented is a method for correlation of the experimentally determined equilibrium constants, together with an example of the use of the method in adjusting these data for application to another gas-condensate system of similar composition. FIELD TESTING AND SAMPLING In order to determine the compositions of the reservoir fluids, field samples were secured from two different wells, one completed in the oil zone, and one completed in the gas cap. The oil well was completed through perforations in 51/2-in. casing at a depth of 6,822-6,828 ft subsea. Pressure and temperature traverses made in the well indicated the pressure and temperature to be 3,831 psig and 201 °F, respectively, at a depth of 6.828 ft subsea. Subsurface samples were taken from this well after a four-hour shut-in period.

Jan 1, 1953

By Ralph B. Stevens

INTRODUCTION Underground fires continue to be one of the most serious hazards to life and property in the mining industry. Although underground mines are analogous to high-rise buildings where persons are isolated from immediate escape or rescue, application of technology to locate and control fire hazards while still in their controllable state is slow to be implemented in underground mines. Even in large surface structures such as hotels, often only fire protection systems which meet minimal laws are implemented due to the high cost of adding extensive extinguishing systems, isolation barriers, alternate ventilation, escape routes and alarm systems. Incomplete and ineffective protection occasionally is evidenced where costs would not seem to be a factor, such as the $211 million MGM Grand Hotel fire November 21, 19801. Paramount in increasing fire safety and decreasing the threat of serious fire is early warning followed by proper decision analysis to perform the correct action. However, very complex fire situations can be produced in structures such as high-rise buildings and underground mines simply because of the distances between the numerous fire-potential locations and fire safe areas. Other complexities arise when normal activities occur that emit products of combustion signaling a fire condition to a sensitive fire/smoke sensor. For example, the operation of diesel equipment or the performance of regular blasting can produce combustion products that reach the sensitive alarm points of many sensors2. Smoke detectors for surface installations provide fire warning when occupants are at a distant location or when sleeping, thus greatly reducing injuries and property damage. However, when installed in the harsh environments of underground mines, fire and smoke detection equipment soon becomes inoperative, unreliable, or requires excessive maintenance. The U.S. Bureau of Mines has performed many studies and tests to improve fire and smoke protection for underground mine workers3. This paper describes several USBM safety programs which included in-mine testing with mine fire and smoke sensors, telemetry and instrumentation to develop recommendations for improving mine fire safety. It is hoped that the technology developed during these programs can be added to other programs to provide the mining industry with the necessary fire safety facts. By recognizing fire potentials and being provided with cost-effective, proven components that will perform reliably under the poor environmental conditions of mining, mine operators can provide protection for their working life and property equal to that which they provide for themselves and their families at home. The basis of this report is two USBM programs for fire protection in metal and nonmetal mines4,5 and one coal program6. The data was collected beginning in May 1974 and continuing through the present with underground tests of a South African fire system installed at Magma Mine in Superior, Arizona, and a computer-assisted, experimental system at Peabody Coal Mine in Pawnee, Illinois. The conduct of each program was as follows: • Define the problem and its magnitude in the industry • Develop concepts to solve or diminish the problem • Review available hardware or systems approaches to fit the concepts • Install and demonstrate the performance of a prototype system through fire tests in an operating mine. MINE FIRE FACTS Whether in coal or metal and nonmetal mines, the potential severity of fire hazard is directly related to location. As shown in Figure 1, fire in intake air at zones A, B, C or D can cause contamined air to route throughout the mine quickly if not detected, isolated or rerouted. Causes and location of former metal and nonmetal fires are represented in Table 1; the cause and location of fatalities and injuries is shown in Table 2. Coal-related fires and their impact on deaths and injuries are graphed in Figure 2; their locations are described in Table 37. Significantly the table shows that the hazard to personnel was three times greater for fires occurring in shaft or slope areas, and the percentage of deaths and injuries was four times that of other areas. Number of Persons Affected A 129-mine sample indicated that from 8 to 479 employees per shift work in underground metal and nonmetal mines, and that deeper mines have larger populations, as shown in Figure 3. Coal mining relates similar employment, and a 16-state sample of 670 mines employing at least 25 persons shows the distribution in Figure 4. Drift mines accounted for 58 percent of the sample but employ only 45 percent of the underground workers.

Jan 1, 1982

By Nicholas Chlumecky

To date information is somewhat limited concerning the strength and bearing capacity of floor strata, which is of great importance where proper roof control is desired. If the bearing strength of strata is not known, the support may be excessively strong in relation to the support capacity of the floor, with the result that strong props sink into the floor strata, giving them dubious value as a roof support. The method described is used to determine how tests are made on dry rock to determine the required support area for the base of props or other structures which may rest on rock strata in mines. Relationships were established for the bearing capacity of rocks which were subjected to loads by means of bearing areas of various sizes ranging from 0.09 sq cm (.014 sq in.) to 700 sq cm (108.5 sq in.). This permits the evaluation of the bearing capacity of rocks by testing small areas for the de- termination of the capacity of larger areas.

Jan 3, 1968

By Arthur F. Johnson

Plant sites for the light metal industry must be located where ample low cost power is available. In the first half of the century hydroelectric development was the only source of this power-now the best sites have been exploited and those remaining do not promise enough water to permit year-round capacity power plant operation. In searching for economic plant sites engineers have concentrated on locations with vast water power potential. Olin Mathieson Chemical Corp. has been studying the economics of developing water power to make aluminum. However natural gas, lignite, and even atomic energy have also been considered as possible power sources for aluminum production. Economic studies have given consideration not merely to investment and production costs, but to total costs of delivering fabricated products to the market center of the U. S. When delivered costs are considered and when full account is taken of the newest techniques of mining bituminous coal and utilizing it to make electric power, one arrives at the inescapable conclusion that coal has emerged as the economic answer to American aluminum production.

Jan 4, 1955

By Clifford J. Lewis, James L. Drobnick

Vexed with competition from relatively high grade foreign orebodies, perplexed by marginal, low grade domestic orebodies, and ironically faced with the fact that practically every metal needed in our economy is available domestically in ample quantity in the ground, our mining industry awaits the advancements in beneficiation technology that will bridge the gap between the ground and the market place for our presently unexploited mineral reserves. Among the vast number of research and development projects underway in the metals mining industry, there are a few chemical techniques in metals winning that, with further development or wider application, could have a profound effect on the industry. These techniques include the Imperial Smelting and Refining Process for complex lead-zinc ores, chlorination and fused salt procedures, the liquid ion exchange process, and the field of non-aqueous leaching. IMPERIAL SMELTING PROCESS The Imperial Smelting and Refining Process is of interest because it offers a possible economic route for the processing of complex zinc-lead-copper ores. The process represents the outcome of the combined efforts of metallurgists and engineers of Imperial Smelting Corporation. Ltd., the United Kingdom Subsidiary of the Consolidated Zinc Corporation, Ltd., and is the subject of patents throughout the world. This process is essentially the employment of the blast furnace for the production of zinc. Zinc has always resisted blast furnace technology, primarily because of the difficulty of condensing zinc vapor from dilute gases. The Imperial Smelting Process employs a blast furnace working at ordinary pressures and in a system which is able to con- dense a blast furnace gas of a very low zinc content.

Jan 11, 1963

By J. L. Leland Myer

Some of the early experiments on the oxidation temperature of coal1 were undertaken in England in connection with a study of self-heating, or spontaneous combustion. It was then suspected that coals with low oxidation temperature were most likely to ignite by self-heating, but further experiments showed that many additional factors were involved. The oxidation temperature has importance as a physical property or characteristic of the coal. While only a few of the physical properties have been studied, they appear to have basic significance2 in the ultimate classification of all coals. When reliable data become available on the oxidation temperatures, as determined by standard and adequate methods, it should provide an additional basis for classification and comparative evaluation. Oxidation Temperature Defined For such a heterogeneous substance as coal is known to be, no simple definition of oxidation temperature can be established on a theoretical basis. If coals were homogeneous and of simpler structure, the accepted definition of "ignition temperature " as applied to simple gaseous mixtures might be adopted; that is, the temperature at which self-heating of the mixture sets in. But most coals self-heat slightly even at storage temperature, as evidenced by spontaneous combustion, and the definition for gases cannot be applied to coals with any exactness. By subjecting coals to suitable experimental conditions various oxidation characteristics can be obtained and examined for critical or significant stages of the oxidation process. Oxidation temperature itself must then be defined on the basis of such experimental results. Experimental Methods Attempts to determine critical points in the temperature functions of coal undergoing oxidation have resulted in the following experimental methods:

Jan 1, 1932

By Edward W. Mehal

This paper presents a review of the technologies which have been used in the application of III-V compound semiconductors to integrated circuits and arrays. These materials have properties which make them useful for various devices such as: light emitters, sensors, microwave sources, and microwave diodes. Technologies for incorporating these devices into integrated circuits have been reported and processes developed for fabricating planar devices using selective epitaxial deposition or selective diffusions. Isolation methods utilizing semi-insulating GaAs, oxide isolation, and ceramic insulation have also been reported. These techniques, and the devices and circuits produced or proposed are discussed. AFTER the invention of the silicon-integrated circuit, methods for utilizing compound semiconductors for integrated circuit applications were proposed. The scope of applications was potentially much broader because of the range of properties available from these materials. There also was the opportunity of coupling these properties with germanium and silicon devices to realize functions and circuits which could not be produced with silicon monolithic circuits. Table I exemplifies the range of electrical properties available in III-V compound semiconductor materials which are currently being widely used. A wide range of sensor and emitter applications result from the range of band gap. The band structure of GaAs and GaAsxP1-x permits their use as microwave oscillator source, and the high electron mobilities further facilitate their utility in microwave devices. The concept of "integrated functions"1 results from the availability of these materials and methods for utilizing their special properties. Heteroepitaxial growth allows the fabrication of a monolithic block of these materials. Semi-insulating GaAs permits the attainment of electrical isolation between components in a monolithic block without resorting to p-n junction biasing or oxide isolation methods. An "integrated function" is a unit which can perform a system or subsystem role and perform a number of electromagnetic functions. These functions could include sensor or input transducer, signal processing, interconnections, and control or output function. Two examples which demonstrate this concept are an optoelectric processor and a microwave receiver front end. The optoelectronic processor, Fig. 1(a), would consist of a light sensor, signal processor, and output emitter. The sensor could be InSb or InAs which would be IR-sensitive. The signal from this sensor would then be amplified and transmitted to the light emitter. The amplifier could utilize GaAs bipolar or field effect transistors. A Gap emitter would then produce visible radiation. The microwave receiver front end, Fig. l(b), would consist of a local oscillator section and a mixer section. A signal input from an antenna is mixed with the signal from the local oscillator. The local oscillator signal is produced by a Gunn oscillator. The single-ended mixer utilizes one Schottky barrier mixer diode. The IF signal would then be amplified further and suitably displayed. A further extension of the capability of utilizing the

Jan 1, 1969