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By R. Schuhmann

Distribution of material in the fine sizes of a comminution product generally is well represented by the empirical equation' y = 100 (x/k)a [1] in which y — cumulative percent finer, x = particle size, a -= distribution modulus, and k = size modulus. Charles3 found that the energy consumption in comminution is usefully expressed by another empirical relation, E = Ak(1-D) [2] in which E = energy input per unit volume of material, A = a constant, k = size modulus based on Eq. 1, and n = a constant; (1-n) is the slope of a plot of log E vs log k. Holmes3 has presented energy equations similar to Eq. 2. The constants a and n in Eqs. 1 and 2 have been shown to depend both on the nature of the material and on the comminuting device. Moreover, Charles showed that within experimental error a and n are a — n+1 = 0 [3] Combining Eqs. 2 and 3, E = A k-a [4] In the first sections of this article it is shown that the energy equation, Eq. 4, can be derived directly from the size distribution equation for the fine sizes, Eq. 1. The derivations are made without assuming any of the specific relationships between energy and particle size which have been common in previous literature. For comminution processes in which Eqs. 1 and 4 adequately represent the experimental data, the constant A in Eq. 4 is found to be a simple and useful inverse measure of grindability. That is, A is the energy consumption per unit volume of comminution product finer than unit size as determined from the straight line portion of the log-log plot of the size distribution. These considerations all lead to a unifying hypothesis of comminution mechanism from which both Eq. 1 and Eq. 4 can be derived. Finally, it is pointed out that this hypothesis raises serious questions as to the significance of the Rittinger hypothesis, the Kick hypothesis, and other theories in which energy numbers are systematically assigned to various size fractions of comminution products in order to calculate theoretical energy consumptions. Derivation of the Energy Equation from the Size Distribution Equation: For simplicity, consider the comminution of 100 volumes of a feed material of relatively uniform particle size. The comminution process may be considered as the summation of many individual and independent comminution events. The extent of comminution is most easily expressed as the number of comminution events, z. In the first derivation, the key assumption is that the characteristics of the comminution events in a given crushing or grinding process are substantially constant and do not vary with the progress of the comminution process. Accordingly the characteristics of an average comminution event may be defined. In one such event, a quantity of energy $E is applied to a single particle of size f and volume $v. The crushing of this particle produces fine particles with a size distribution similar to that given by Eq. 1: yw=100(x/ka)a0 [51 In this equation yo, a0, and k0 are used to characterize the product of an individual comminution event rather than the product of the comminution process as a whole. In using Eq. 5, we will not be concerned with values of x close to the feed size f and will therefore assume only that the equation is applicable to the finest sizes of the material. In 100 volumes of total product, the actual volume of product finer than x from a single comminution event, or dy, is given by The total volume of material below size x, resulting from z events, is then given by y = z(dy) =z(dv) (x/ka)a0 =" Eq. 7 reduces to Eq. 1 when we let a, = a and [8a] z =100/dv (ka/k )1 or k = ka (zdv/100)-1/a [8b] Eq. 8a shows that the distribution modulus of the comminution product is the same as for the product of an individual comminution event. Eq. 8b shows how the size modulus of the comminution product k varies with the extent of comminution as measured by the number of events, z, or as measured by the fraction of the feed actually subjected to com- minution1 zdv/100. The energy input to 100 volumes of total feed, or 100E, is the sum of the energy inputs for all the comminution events: 100E =z (dE) [9]

Jan 1, 1961

By Charles Will Wright

THE economy as well as the living standards of a country depends largely upon adequate supplies of raw materials at reasonable prices. Geological and climatic conditions responsible for the occurrence of commercial minerals and the growth of raw agricultural and forest products are so varied in different parts of the world that no country can be self-sufficient. The world is rapidly shrinking because of improvement in transportation facilities, but populations are steadily increasing and so are the needs for a greater production of raw materials and for their more equitable distribution. Too many barriers between nations prevent free access to the main sources of supply of the raw materials vital to neighboring nations and it is just such barriers that are often the cause of friction and even war. The International Trade Organization is doing a commendable job in helping to. break down barriers and to reduce friction, and we should now find ways to facilitate access to these sources of supply and encourage more efficient and greater production of the raw materials so urgently needed throughout the world. This cannot be done by individual selfish governmental agencies, influenced as they are by politics, nor by socialistic governments attempting to nationalize industries. It can be done efficiently only by the free individual through private enterprise properly protected against adverse governmental action. This protection can be supplied by the World Bank.

Jan 1, 1948

By John B. Knaebel, Robert J. Grant

This paper outlines the trends in gold production since the discovery of America, in the world as a whole, and in the principal producing regions as well. World production climbed at an average rate of about 20,000,000 oz. per 5-year period until the record all-time output of 111,307,000 oz. in the period ending with 1915. Production fell off during the war period, but has been mounting since 1925, and will probably be nearly 100,000,000 oz. for the 5-year period ending in 1930. Total world production of gold from 1492 to the end of 1930, as closely as can be determined, is more than 1,061,000,000 fine ounces. More than one-half of this total, or about 575,000,000 oz., has been mined since 1900. In percentage of world production, the United States jumped from 4.8 per cent for the decade 1831-1840 to 44.5 per cent for the five years ending with 1855. This is the highest figure ever reached by this country, although, with the exception of the 10-year period ending 1865, it ranked first among the nationa from 1849 until the end of 1905. During the 5-year period ending with that year, the United States produced 24.7 per cent of the world's gold. The tremendous increase from South Africa following 1890 resulted in that country's becoming the largest contributor for the five years 1906-1910, a position it has held with increasing dominance ever since. Relegated to second place in the period 1906 to 1910, the United States produced nearly 20 per cent of the world's new gold until 1920. From 1921 to 1925 it averaged 13.6 per cent and since then it has continued to decline in relative importance, accounting for only 10.9 per cent in 1927. Final figures for 1930 may place Canada second among the gold producers of the world.

Jan 1, 1931

By Bezalel C. Haimson

We have developed a new wireline hydrofracturing technique for in situ stress measurement which eliminates the need for drill rig and drill rod and renders the tripping of the straddle or impression packers continuous, fast, and inexpensive. Wireline conductors facilitate the use of downhole pressure transducers. We have confirmed the suitability of the new technique by testing it in a deep hole near the Monticello Reservoir, South Carolina. Our measurements indicate two local stress fields, one of them apparently linked to the impoundment-related induced seismicity.

Jan 1, 1984

By K. L. Young

The general energy equation, including change in kinetic energy, was solved by numerical integration and used to evaluate simplifying assumptions and application practices over a wide range of conditions. When extreme conditions were encountered, sizable errors were caused by large integration intervals, application of Simpson's rule and neglecting change in kinetic energy. A maximum error of only 1.31 percent was caused by assuming temperature and compressibility constants at their average value. It was discovered that a discontinuity can develop in the integral for the injection case. This discontinuity indicates a point of zero pressure change and is an inflection point in the pressure traverse. INTRODUCTION When a pressure in a gas well is to be calculated, one of the first decisions is to select a method of calculation. In many instances, this selection becomes a problem because the literature, at best, provides an evaluation of any method for only a limited range of conditions. Once a method has been selected, a question often arises as to the size of the calculation interval which should be used. The question regarding calculation interval arises because an analytic solution is not obtainable and approximate solutions must be used. This paper presents an evaluation of major assumptions and application practices of probably the two most widely used methods for calculating steady-state single-phase gas well pressures. The two methods are Cullender and Smith1-" (numerical integration), and average temperature and com-pressibility. The Cullender-Smith method assumes that change in kinetic energy is negligible and is normally applied in two steps with a Simpson's rule correcton. The average temperature and compressibility method, in addition to neglecting kinetic energy change, asumes that temperature and compressibility are constant at their average values. This method is normally applied for wellhead shut-in pressures of less than 2,000 psi, and in one step. Computer programs were written to compute bottom-hole pressure with and without the assumptions, using various approaches. Values of input parameters investigated are shown in Table 1. Flow rate was limited to a maximum of 5,000 and 10,000 Mcf/D for tubing sizes of 1.610 and 1.995 in. ID, respectively. Flow rate was also limited to 10,000 Mcf/D for a tubing size of 2.441 in. ID when wellhead flowing pressure was 100 psia. These limitations were imposed on flow rate so as not to exceed sonic velocity. The z factor routine available necessitated limiting bottom-hole temperature to 240F and wellhead pressure to 3,000 psia. Pressures were compared on the basis of percent deviation from the trapezoidal integration of Eq. l or 2 at 100-ft intervals. A preliminary investigation indicated that a 1,000-ft interval solution would differ from a 50-ft interval solution by less than 0.25 percent; therefore, the 100-ft interval was chosen for a base. For the purpose of comparison, deviations less than 1 percent were considered insignificant. where 111.1 qeC2/d p = kinetic energy term. Eqs. 1 and 2 can be evaluated numerically at specific depths using the trapezoidal rule as shown by Cullender and Smith. If change in kinetic energy is neglected and temperature and compressibility are assumed constant at their average values, Eq. 1 can be integrated to give the average temperature and compressibility equation,

By D. Havlena, A. S. Odeh

The use of the straight-line method of solving the material balance equation is illustrated by means of six field cases. Also, the application of statistical criteria to arrive at the most probable answer is shown. The theory underlying the straight-line method of solution and the applicability of the statistical criteria was presented in a previous paper.' The field cases include saturated and undersaturated oil reservoirs with and without water drive. The aquifers discussed are: limited radial, infinite radial, very small aquifer and infinite linear. The field cases also include a gas reservoir producing under water drive. INTRODUCTION In a previous paper,' the authors presented the theory underlying the solution of the material balance as an equation of a straight line. The appropriate equations for various material balance cases as well as the methods of analysis and interpretation with comments and discussion were also included. To illustrate the various theoretical cases treated previously, five field cases are analyzed in this paper by employing the straight-line method of solving the material balance equation (MBE) and one example previously published is referred to. The use of statistical criteria to arrive at the most probable answer is also shown. All the field examples presented, except Case 2, are excerpts from complete reservoir studies. To illustrate the method, only sections specifically dealing with the material balance principles are included. Additional geologic information and basic data are reported to better acquire an understanding of the cases and thus to better follow the reasoning that suggested the successful application of the straight-line method of solving the MBE. The six cases are: (1) saturated reservoir, small gas cap, limited aquifer; (2) saturated reservoir, very small gas cap, infinite aquifer; (3) undersaturated-saturated reservoir, very small aquifer; (4) highly undersaturated reservoir, no water drive; (5) high undersaturated one-well reservoir, limited aquifer; and (6) gas reservoir, infinite linear aquifer. THE D4 SAND, GUICO FIELD, VENEZUELA The D, sand, which was discovered in 1943, is presently in a depleted state. Since its discovery it has produced under water drive, gas-cap-gas expansion, and solution gas drive. In Nov., 1947, water injection was initiated to arrest further pressure decline. When discovered, the D, sand was a saturated reservoir with a gas cap /oil zone volume ratio m estimated volu-metrically at 0.0731, an average permeability of 500 md, a porosity value of 25 per cent, and an oil viscosity at reservoir conditions of 0.3 cp. The volumetrically determined stock-tank oil initially in place was 23.1 million bbl. The volumetrically weighted physical data and production data available until Nov., 1953 are reported in Table 1. In Ref. 1, the effects on the straight-line plot of various values of ro/r for a constant ?to, or of various dimen-sionless times for a constant r8/r10 were theorized and were illustrated in Fig. 3A of that reference. In this field case, the previously theoretically predicted effects are established. Thus, the MBE calculations using Eq. 3c of Ref. 1 were performed for various r6/r10 and dimension-less time values. Eq. 3c of Ref. 1 is:

Jan 1, 1965

By Henry S. Rawdon

The structure of wrought iron as usually described by metallographists and workers in metal in general is that of a fairly pure iron. Impurities, if present, are usually considered as being in solid solution in the crystals of the ferrite matrix or as forming part of the ever present "slag streaks." Attention is herein directed to one type of these dissolved impurities; to the detection of such impurities, particularly in low-grade irons, and to their possible influence on the physical properties of the metal. The attention of the Bureau of Standards was first directed to wrought iron of the peculiar and unusual characteristics to be described later, in material which had failed in service and was submitted for test. The examination was extended to other grades of wrought iron, to see whether such features are of common occurrence in this class of material. MICROSTRUCTURE OF WROUGHT IRON Usual Structure The general microstructure of wrought iron is so well known that a detailed description of it here is needless. There is shown in Fig. 2 the structure of a sample of Swedish iron which illustrates well the matrix of ferrite crystals in which are embedded the slag threads so characteristic of the puddling process and the subsequent working of the material. The usual commercial product contains varying amounts of the impurities commonly associated with iron and steel, i,e,, manganese, phosphorus, silicon, and sulphur together with small amounts of carbon. Of these, the larger part of the phosphorus and some of the silicon present are held in solid solution in the ferrite while most of the remaining impurities, other than carbon, exist in the in closures of slag. Aside from the discontinuities introduced by the slag streaks the ferrite matrix has the microstructure and appearance of a pure metal. No definite orientation of the various crystals or grains is apparent; deep etching of the specimens reveals no intracrystalline features (i.e., within the grains) other than those noted in " pure " metals in general (etching pits).

Jan 1, 1918

By Hans Conrad, W. D. Robertson

FORMAL theories of yielding and flow have been developed to describe thermally activated plastic flow in metal crystals.'-" hey are all based on Becker's' original idea that thermal fluctuations aid the applied stress in overcoming internal barriers to flow and lead to the following relation. where j is the strain rate; C is a constant which includes a frequency factor, the entropy, the number of flow units, and the amount of strain associated with each successful fluctuation; E(T) is the activation energy associated with flow, and is a decreasing function of the applied stress, 7; k is Boltzmann's constant; and T is the temperature in OK. The various theories differ mainly in the model chosen and in the nature of the relationship between the applied stress and the activation energy. Beckerl and Orowan' suggest a relationship of the form E a(Ti — t)2, where .ri represents the stress required for flow without the aid of thermal fluctuations; Kauzmann,3 Cottrel1,5 and Seeger6,7 suggest E a(t — t) ; Fisher8 suggests E while Cottrell and Bilby4 suggest a more complex relationship between E and 7. Additional data are needed to test these various theories., esweciallv- in view of Dorn's9 ork. which indicates that the activation energy for creep is independent of the stress and strain. The present paper presents the stress-strain data derived from incremental loading of magnesium single crystals in tension and direct shear, in the temperature range of 78" to 364°K. Incremental loading was employed in order that the creep component of the strain could be determined. Stress-strain curves of aluminum10-13 and copper" single crvstals. derived from incremental loading tests, have been found to be in agreement with those obtained by the more conventional constant strain rate or loading rate tests. Two types of tests were conducted in the present investigation: 1) tests in which the temperature was maintained constant throughout the entire

Jan 1, 1958

By George Anagnostache

AMONG the petroleum-producing countries of the world, Roumania occupies the fifth place; in Europe, it occupies second place, coming after Russia. The extraction and the use of petroleum in Roumania, may be trace( as far back as 1856, but the production on an industrial scale started in 1896, when the "Steaua Romana," was formed. From that time until the beginning of the war, the petroleum industry had a rapid growth the number of the producing companies increased to 37, and production FIG. 1.-ROUMANIAN OIL AND GAS FIELDS. increased from 81,571 tons, in 1896, to 1,898,545 tons, in 1912, when the greatest production was reached. In 1916, when the Germans invaded Roumania, the oilfields, at the order of the Allies and under British supervision, were destroyed, to pre vent their exploitation by the Central Powers; the wells were blown in the refineries and the storage tanks burnt, and pipelines put out of order This voluntary destruction, which weakened the resistance of the Central Powers, set back the petroleum industry at least one decade. Reconstruction after the war was slow, because of bad transportation conditions the scarcity of drilling and refining material, the insufficiency of. electric power, and the continuous depreciation of the Roumanian money, which made difficult the buying of the necessary materials for reconstruction.

Jan 3, 1925

By W. G. Fischer

This paper attempts to close the gap between extensive rock testing with field measurements and the mine operator's need for safe mine layouts and alternatives. It was found that many of the variables that influence mine geometry and worker safety are relatively unimportant while others are critical. The development of empirical relationships that tie the critical variables to the safe useful life of a trona mine layout and observations made in 140 production areas where the variables were researched over a 15-year period are discussed. The concepts presented may have applicability in mine design for other bedded deposits.

Jan 1, 1986

By J. A. Knox, R. M. Lasater, J. M. Tinsley

Chemical squeeze treatments have been used to provide temporary relief from certain production problems. The chemical fracture-squeeze technique, combining the effects of a fracturing treatment and a squeeze operation, has been more successful than conventional squeeze operations. Knowledge derived from well stimulation and reservoir engineering research provides a means for predicting the theoretical effective life of such a treatment. Analysis of theoretical equations and concepts developed allows selection of improved treatment techniques based on specific formation conditioins. Theory used in this analysis was developed as an extension of previous electrical model studies made to establish the expected flow and pressure profiles adjacent to a fracture system. The chemical fracture squeeze technique can be utilized in the economic application of corrosion inhibitors, emulsion breakers and paraffin and scale inhibitors. Application of this technique is shown to be effective. The slow return rate of injected chemicals, controlled by the resultant flow profiles and treatment variables, permits extended periods of chemical effectiveness. Results of field treatments are given, showing that the concepts outlined above for chemical fracture-squeeze treatments are valid and that applying this technique can help alleviate many current production problems. INTRODUCTION Much progress has been made in the last 10 to 15 years in developing chemicals for use in stimulating wells, maintaining production and protecting well equipment from damage due to corrosion. Not too many years ago, some wells seemed to dry up or wear out. In many cases the wells were produced as long as possible without any attempt at maintaining productivity. Even with the advent of new and better stimulation techniques, a rapid decline in production was observed. Methods of removing and, in some instances, preventing damage have been developed. Among thosc factors responsible for uneconomical production are scale, paraffin, corrosion, bacteria, water blocks and emulsions. Soluble scale-prevention chemicals have been developed1,2 that can be placed in a formation along with frac- turing sand. As the water produces back across this bed, the solid material dissolves slowly and can provide long-term protection from scale. However, bottom-hole temperature and salinity of produced water vary widely and both these factors influence the rate of solubility. Scale inhibitor composition is also a controlling factor. Some of the solid material may be crushed, increasing the surface area exposed to water and increasing the rate at which it dissolves. Some of the material may never be contacted by water and can be lost. However, this type of treatment has been very successful in many instances and has helped maintain economical production for extended periods of time. Liquid scale inhibitors, which are more widely applicable and more stable, have been developed in recent years; however, because they are liquids, their use has been restricted to treatment down the annulus, using metering pumps to provide proper concentrations in the produced fluid. This has prevented use in wells containing packers, in dually completed wells and in gas-lift and flowing wells. Wells that operate with an open annulus may also experience severe corrosion problems due to introduction of oxygen. Paraffin inhibitors3 have been developed that can be fractured into a well as particulate solids to be slowly dissolved in the produced fluid. These materials are not usually effective in wells with a bottom-hole temperature in excess of 120F since solubility rate may be too fast if that temperature is exceeded or if aromatic content of the oils is unusually high. Corrosion inhibitors have been developed that can be fractured' into a well for long-term feedback, but development of a material with proper solubility or feed rate has been difficult. Corrosion inhibitors are available in many different forms. Liquids have been lubricated down the annulus or sticks or pellets dropped down tubing. Inhibitor squeeze treatments5 devcloped a few years ago led to development of inhibitors with particularly strong film-forming properties.6,7 This technique basically involves displacing a highly concentrated solution of the inhibitor into the formation through the tubing. Kerver and Hanson8 studied the adsorption properties of inhibitors on various types of formations. They showed that, even though the inhibitor was displaced radially into the true permeability, it could be produced back for a long period of time because of slow desorption from the rock. Methods developed for monitoring the return of these inhibitors generally have established 1 to 6 months as the effective limit before retreatment is necessary.9 Inhibitors displaced into the interstices of the formation sometimes cause emulsions that either hamper production or cause treating problems on the surface.

By P. D. Rao, H. B. Charmbury, D. R. Mitchell

As a result of a changing market pattern for anthracite, there is now an increasing demand for steam size and a decreasing demand for domestic sizes. To help the producer meet this new demand, the authors present the technical aspects of the problem of crushing domestic sizes to steam fineness and size. The market pattern for anthracite coal has changed considerably in recent years. The demand for domestic sizes has been gradually decreasing and the demand for the steam sizes has been increasing. The companies however aim for a maximum production of domestic sizes, since these sizes command premium prices. To meet the market demands, particularly in certain periods of the year, further size reduction of the coarse clean coal is necessary. The present practice of multistage crushing in toothed rolls leads to complicated flow arrangements and is high in cost. Data presented herein indicate that crushers with high reduction ratios could be used with a resulting simplification of crushing flows and an appreciable reduction of capital and operating costs. The different types of crushing equipment investigated included the rotary-type breaker, the jaw crusher, the impactor-type crusher, and the Aerofall mill. CURRENT PRACTICE The standard sizes of anthracite are presented in Table I and market specifications require that any given size be maintained with only a limited amount of oversize and undersize material. The 1958 prices of the various sizes also are presented, so it may be seen that size reduction is only performed when dictated by market demands. At the present time the general practice of crushing anthracite consists of removing the top size and crushing it in a toothed double roll crusher to the next largest size. Consequently, to reduce the domestic egg, stove, and nut sizes to steam sizes of pea, buck, rice, and barley, the original material must be screened to produce the three top sizes and each size must be crushed in three separate rolls to the next largest size. Each crushed product must then be rescreened to remove the oversize for re crush ing. Thus a typical operation would consist of scalping the egg size from a mixture, crushing it to stove size, and rescreening the product. This same cycle would be repeated for the stove and nut sizes requiring three screens and three crushers as well as conveying equipment to handle the crushed products. This results in an expensive series of operations to make a product which sells for a lower price. If the three top sizes are crushed in one operation to the smaller sizes then the yield of steam sizes is low and the production of Nos. 4 and 5 sizes is large. These sizes sell for lower prices than the steam sizes. A typical operation is presented in Fig. 1. From the figure it may be seen that the original cleaned anthracite contains 67.4 pet domestic sizes, 26.4 pct steam sizes, and 6.2 pct fines. After multistage crushing of the domestic to steam sizes the composition of the product is 0.6 pct domestic sizes, 89.6 pct steam sizes, and 9.8 pct fines. From a technical standpoint these results are reasonable but the handling and crushing costs are

Jan 1, 1961

By Morgan Walker

Intermittent injection of gas in gas-lift pumping is a variation of the common practice in that the gas is delivered to the well for a short period, called the "on time," followed by a period during which none is admitted, called the "off time," instead of being admitted continuously at a constant rate or by intermittent flowing in which it is occasionally delivered to the well in just sufficient quantity to flow off the accumulated head of fluid. The term "gas" is used throughout this discussion but it should be remembered that in so far as the operation is concerned either air or natural gas may be used. The equipment at the well is the same as commonly used in gas-lift installations and gas may be admitted to either the tubing or the casing. It seldom occurs that all the wells on a lease can be pumped economically by intermittent injection from one header. Wells that have flow characteristics which permit their being handled together have their intake lines grouped at a central compressor station. Well groups may consist of two or more wells. The total volume of gas delivered by each group of compressors goes to one well at a time through a quick opening valve, each well connected to the header receiving its charge of gas in turn. Special cases may require a deviation from this arrangement. The distribution of the gas to the different wells is effected by an automatic timing device which operates quick-opening valves on the intake lines. The timing mechanism may be actuated by electricity, liquid or gas pressure, The essential factor in the operation of the timing mechanism is that as one intake valve is closed another must be opened. Water-operated Timing Device Timing devices operate the intake valves either directly, or indirectly by pilot valves. Of the former type one that is operated by water and has given satisfactory service is shown in Fig. 1 and can be made on the lease. The horizontal cylinder is a piece of 10-in. casing about 7 ft. long, closed at each end and free to oscillate about a fulcrum at the center.

Jan 1, 1929

By E. T. Turkdogan, P. Grieveson

Experimental results are given for the rate 0.f solution of nitrogen in y iron in the temperature range 1000° to 1200°C. It is shown that, when purified reacting gas is used, the rate-controlling process is the diffusion of nitrogen into the iron. The difhsiuity of nitrogen in y iron is calculated from the results and the equation logD = —(8800/T) — 0.04 is recommended to represent the temperature dependence of the diffusivity. In some experiments using unpurified nitrogen, solution rates are found to be slower than those attributable to a solely solid-state diffusion process. Indications are that in the presence of impurities in the gas, e.g., H,O, chemisorption of nitrogen at the surface is hindered and becomes a rate-controlling process for the transfer of nitrogen across the gas -metal interface. ALTHOUGH the solubility of nitrogen in y iron at known nitrogen activities has been well-established,'-a the kinetics of the solution of nitrogen in iron are not known. The work of Darken, Smith, and Filer' indicates that at temperatures below about 1300°C the rate of solution of nitrogen in y iron is controlled by a chemical reaction at the metal surface, while at temperatures above about 1300°C the rate-controlling process is the diffusion of nitrogen in the metal. Fast and verrijps studied the denitro-genization of y iron using wet hydrogen at 950°C and their results clearly indicate that the rate of nitrogen removal is controlled by diffusion of nitrogen in the metal. In Part I of this series of papers, the kinetics of the solution of nitrogen in y iron are discussed in the light of the present experimental work. EXPERIMENTAL Apparatus. The reaction chamber consisted of a platinum wire-wound resistance furnace with a re-crystallized alumina reaction tube. The temoerature of the furnace, which was automatically controlled to approximately 1°C, was measured with a Pt/Pt-13 pct Rh thermocouple and a potentiometer. The temperature distribution in the working zone of the furnace was determined using a calibrated Pt/Pt-13 pct Rh thermocouple and the hot zone was found to be approximately 3 in. long with a temperature variation of * 3°C. A series of experiments was also carried out at 1000°C and pressures above atmospheric using a pressure vessel which was similar in construction to that described elsewhere.1° Materials. Ferrovac E grade iron used in all ex-

Jan 1, 1964

By G. B. Misra

While investigating on the aerodynamic resistance of airways supported with peripheral timber sets, at regular intervals, the following theoretical equations were developed by the author to estimate the resistance coefficient of such airways: [ ] for S < 1, where f is Darcy-Weisbach resistance coefficient of the airway, C is modified drag coefficient of the supporting member, D is equivalent diameter of the bare airway, 8 is ratio of the approach velocity over the sets to the average velocity of the bare airway, A is cross-sectional area of the bare airway, a is projected frontal area of the sets, A., is cross-sectional area of the air stream at the vena contracta inside the set, S is spacing of the sets, f, is resistance coefficient of the bare airway, l is length of aerodynamic influence of sets, p is perimeter of the bare airway, p, is setted portion of the perimeter of the bare airway, pe is unsetted portion of the perimeter of the bare airway, and P shielding factor. The equations were verified experimentally in a model rectangular airway supported with one- (bars), three-, and four-piece sets of square-section timber of three different sizes and were found to hold true. The work has been further extended to one-, three-, and four-piece sets of round timber of 2.6, 3.2, and 3.8 cm diam with the same experimental set up. Tests have been carried out for spacings of 25, 50, 75, 100, 150, and 200 cm over a regime of flow defined by the Reynolds number (with respect to the equivalent diameter of the bare duct) ranging from about 1.5 X 106 to 5 X 106 using the same experimental techniques. The values of f are calculated in the manner indicated in [Ref. 1]. Unlike with square-section timber, the resistance coefficient f of the airway setted with round timber shows a distinct variation with the Reynolds number of flow. This conforms to observations made by Sales and Hinsley.2 In order to have a comparable value of f for all types of sets with all sizes of timber, it was necessary to select the value of f at a fixed Reynolds number of flow. Since f is chiefly a function of the drag coefficient of the sets, the appropriate Reynolds number RE is that with respect to the diameter of timber in the set. Considering the diameters of timber used and the regime of flow over which measurements were made, f was chosen at a value of RE = 20,000 in all cases. The f vs. S curves are maximal in nature and in conformity with theory, the f vs. 1/S curves are straight lines up to a value of S = 1 beyond which they show a distinct flexure. The observed values of 1, the length of aerodynamic influence of sets, agree with the relation 1 = 42 e, developed for square-section timber sets, thus suggesting that the shape of timber has little influence on the length of aerodynamic influence. The value of the modified drag coefficient CD for round timber was calculated in the same way as for square timber in Ref. 1, taking the contraction factor Z = 1.5 for round-edged constrictions. CD has an average value of 0.96 with a standard deviation of 6.08% as compared to the free stream drag coefficient of 1.2 at RE = 20,000 for long cylindrical obstructions The shielding factor [ ] is plotted against S/1 in [Fig. 1]. The curves are more or less independent of the size of timber, but are different for the different types of sets, possibly due to their different degree of symmetry. Values of f calculated by the author&apos;s [Eqs. 1 and 2], using experimental values of CD&apos; and [ ] and taking I = 42 e, are plotted in [Fig. 2] against experimentally measured values of f for different types of sets with different sizes of round timber. The values agree closely with a standard deviation of only 5%, thus establishing the veracity of the theoretical equations developed by the author for round timber as well. A comparison was made between the Xenofontowa4 equations (the only other reasonable relations available for the estimation of the resistance coefficient of supported airways) and the author&apos;s [Eqs. 1 and 2] by comparing in [Fig. 3] the values of the resistance coefficient f computed by the Xenofontowa relations with those experimentally measured by the author. In order to make

Jan 1, 1975

By Tell Ertl

An annual invitational address by an outstanding man in mining, geology, or geophysics who has contributed significantly to the progress of technology in these fields. IN 1954, the first year the award was conferred, Fred Searles, Jr., remarked, "The earlier years are the easiest." In this, the second year, it was easy to choose the outstanding mining engineer. Our conferee this year has been destined to mine. He comes from pioneer stock, and though a son of the American Revolution he is most proud that all four of his grandparents came to the Jordan River Valley of Utah the "first summer," which all Utahns will recognize as 1847. At the time of discovery of high-grade ore in Bingham, his grandfather, Archibald Gardner, had a crew of men getting out saw timber in the area. The first claim staked in Bingham Canyon was named the Jordan, and Bishop Gardner was the colocator of the claim as well as the man to record it. E. D. Gardner was born in West Jordan, Utah, in view of Little Cottonwood Canyon, the site of the Emma Mine. Park City in the Wasatch Mountains a few dozen miles to the east, and Tintic, not much further to the south, were developing their vast ore-bodies throughout his youth. As a boy he was near the present Midvale metallurgical area and could see the Lark and Bingham Canyon mine dumps and imagine the mining activities at Tooele, Ophir, and Mercur just over the nearby Oquirrh Range. He was 11 years old when the first car of copper ore was shipped from Bingham Canyon. It was not surprising that he attended the University of Utah and graduated with an engineer of mines degree. Five years with the Utah Apex in Bingham Canyon as chief engineer and assistant superintendent completed his training for the 44 years of unselfish and lightly rewarded service he was to give the mining industry as a Government engineer. All of us in mining are familiar with the great work he carried on and directed during the genera- tion he supervised the Southwest Experimental Station at Tucson. He has written scores of papers and books. His observations and experiments on drilling and blasting have led to the important improvements in this field during the past quarter century. Whereas those of us who consider ourselves widely informed about underground metal mining methods make hasty trips through a half dozen or dozen mines a year, our conferee has visited them all—not only the metal mines, but also coal and hundreds of nonmetallic mines, in the United States, Canada, Latin America, and overseas. His keen observations and analytical mind have resulted in an encyclopedic knowledge of mining and mines. His improvements in mining technology have been steady and predictable. His elucidation of mining problems to the many young engineers that have come under his guidance have made them grow strong and useful to his industry. When he chooses to stop, his influence will continue to expand and prosper because of the inspiration and dedication he instilled in so many of us. The profits from his ideas have been distributed throughout the mining industry. He has remained a technical man who felt that he could serve the mining industry best as a Government engineer. But he is the staunchest advocate of the belief that the public should gain from his work through private industry. This is not the first recognition E. D. Gardner has received. The Department of the Interior awarded him the coveted Medal for Distinguished Service in October 1949, and he received the Engineering & Mining Journal&apos;s Salute in March 1951. His friends who are presenting him the Daniel C. Jackling Award today know this great mining engineer simply as "Ed" Gardner.

Jan 1, 1956

The Institute is one of the five national societies that joined in organizing the Congress, the others being the American Society of Civil Engineers, the American Society of Mechanical Engineers, the American Institute of Electrical Engineers, and the Society of Naval Architects and Marine Engineers. Each society has pledged from its general funds or some special fund a certain sum to guarantee the success of the Congress. To this the engineers resident on the Pacific Coast added $13,000, the total guarantee fund amounting to $37,000. It is hoped that this money, iii part or in whole, can be repaid to the guarantors of the Congress. The sums necessary to conduct the latter and to publish the proceedings, are to be derived from membership fees paid equally by all participating. A member of the Institute as such is not a member of the Congress and will not receive any of the publications of the latter. He is eligible to become a member of the Congress and it is hoped that he will do so. The membership fee is $5, and this entitles one to receive the introductory volume of the Congress containing the proceedings; a digest of the other volumes and a general index. It also entitles the member to one other volume of the transactions, to be chosen as he will. For moderate sums he can buy additional volumes, the whole set costing delivered, $30 or less. On Jan. 12 the total membership was 2,350, of which only 290 were recorded as mining engineers. Full details regarding the Congress have been sent to every member of the Institute, and the failure of the members to take a more active interest in the matter is believed to be due to some misunderstanding of the situation.

Jan 4, 1915

By Walter E. Lewis

Myriad problems face all of us in the next decade. Vietnam, poverty, and pollution are perhaps the most pres- sing. A lesser one but still vital to us as a Nation is the slow hut apparently relentless extinction of the mining, metallurgical. geological and petroleum engineer. Are we doomed to take our place in the crossword puzzle with the moa. to be defined like it as an "extinct bird"? Or can we reverse the trend perhaps, in the next decade? The plain fact is we are an endangered species, and must resign ourselves to virtual extinction in the 1970&apos;s unless the trend is reversed. Probably, many of us have felt the problem will solve itself if left in the in-basket long enough, believing, that the trend will reverse solely on the basis of demand for these disciplines. This approach has not worked up to now, it gives no promise of working in the future, and furthermore, the problem has been in the in-basket much too long. The point is simply this: What can each of us do as an individual, or as part of a company, a State agency, a Federal agency, a university, or an organization such as AIME?

Jan 1, 1970

By Dr. O’Neil Thomas J., Donald W. Gentry

It was a western truism that more money was made from selling mines than from buying them, just as it was accepted that many a good mine had been spoiled by working it. from R. E. White, "The Mining Town," Alta California August 18,1873 INTRODUCTION One of the most important aspects of mineral enterprises of US corporations is the influence of the depletion allowance on cash flow. This allowance permits mineral producers to claim sizable federal income tax deductions not available to other industries, and these deductions significantly increase cash flows in mining. In the past several years the minerals depletion allowance has come under increasingly heavy fire from many sectors as a tax loophole that deprives the federal government of large amounts of revenue. One estimate of the loss of tax dollars in 1968 was $2.25 billion due to depletion deductions and intangible expenses in the oil industry. Largely on the basis of such arguments, the Tax Reform Act of 1969 reduced percentage depletion rates for a number of minerals, and the Tax Reduction Act of 1975 totally eliminated statutory depletion for most oil and gas. In regard to percentage depletion for hard minerals, the federal government has estimated its resulting loss of revenue at $1.75 billion for 1980. The TEFRA legislation of 1982 further reduced the statutory depletion allowance available to iron ore, coal, and lignite beginning in 1984. Thus, it is clear that minerals engineers and managers should have a thorough knowledge of the origin, evolution, justification, and computational procedures relating to the depletion allowance. ORIGIN AND EVOLUTION OF THE DEPLETION ALLOWANCE

Jan 1, 1984

By AIME AIME

MODERN metallurgy is an art and a science. The art is process metallurgy-extracting metals from their ores, refining them, and alloying them with one another and with certain nonmetals to produce thereby a variety of metallic materials which, owing to a wide range of useful properties, are for many purposes more valuable to man than the individual metals themselves; process metallurgy also includes shaping or otherwise treating metals and alloys to form objects useful to a complex industrial world. The science is physical metallurgy-organized knowledge of the constitution and internal structure of metals and alloys, of the relation of their composition and structure to properties, and of the laws, theories, and hypotheses that purport to explain this relation.

Jan 1, 1942