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Mr. John J. Porter, Chicago Ill. (communication to the Secretary†):—Mr. Baker's account of his experiences with stock-distribution has been particularly interesting to me, as it

Jan 1, 1905

By H. Vincent Wallace

ALTHOUGH it is universally accepted that free oxygen is A necessary in a cyanide solution for the dissolution of gold and silver-in accordance with Elsner's equation that 2Au + 4KCN + 0 + H20 = 2KAu (CN), + 2KOH- little has been written and insufficient attention paid to the drop in extraction due apparently to loss of dissolved oxygen in mill solutions. Aeration or oxygenation of cyanide solutions has been more or less successfully accomplished by various types of agitators such as the Dorr (with or without air lifts), Pachuca tanks and more recently "Turbo-mixers" and Denver "Super- agitators"; also by the introduction of chemical oxidizing agents such as Solozone (Na,O,) . At the Wright-Hargreaves and other mines in the Kirkland Lake camp, the introduction of this reagent at the rate of about 1.25 oz. per ton of ore applied to the feed of the ball mills has certainly been beneficial.

Jan 1, 1932

By Clarence Zener

THE heat-treatment of steels will not pass from the stage of an art into that of a science until the mechanism of the phase transformations associated therewith is thoroughly understood. Such an under- standing will involve two distinct types of inquiry: (I) the direction of the trans- formation, (2) the speed of the transformation. The first type of inquiry, which involves only the equilibrium relations between the various phases in steel, is the subject of the present and of a succeeding article; the second type, which involves the kinetics of transformations, will be the subject of a later article. In order that a theory of equilibrium relations may give useful information, certain experimental data must be put into the theory. It is customary for these data to be of a thermal nature. In the system discussed here in steel such thermal data are very scarce. In medium-alloy steels another type of data might possibly be used. In these steels the concentration of solute is so low that it may be possible to use the approximation of dilute solutions, an approximation that leads to linear relationships. The unknown functions, or constants, may then be determined in some cases from the binary iron- alloy system and in other cases from the ternary iron-carbon-alloy system. From the functions, or constants, so determined may then be computed the equilibrium relations in more complex medium-alloy steels. In the present article this second approach is used, the necessary experimental data being obtained from the simple binary and ternary equilibrium diagrams. FUNDAMENTAL EQUATIONS In this article are derived the equilibrium relations between the gamma, alpha and cementite phases. In these three phases all important alloying elements except carbon form substitutional solutions, each alloying atom occupying a potential site for an iron atom. Carbon, on the other hand, forms an interstitial solution. In the face-centered cubic gamma phase, the interstitial positions are at the center of the unit cube and at the centers of the cube edges. There are therefore four interstitial positions per unit cell, or one per iron atom. In the body- centered cubic alpha phase, the interstitial positions are probably at the face centers, and also at the centers of the edges, which are crystallographically equivalent to the face centers. There are therefore six interstitial positions per unit cube, or three per iron atom. In cementite, (Fe, Alloy)3C, the carbon atoms may be regarded as filling all the lattice positions of a given type. The introduction of the fundamental equations of equilibrium has presented the author with a quandary. These equations are formally so similar to other equations that have been derived1,2 for related systems that they may appear to some readers

Jan 1, 1946

By Wm. E. Milligan

DESPITE the zero weather of Monday, the morning meeting on nonferrous ore-reduction metallurgy got under way promptly under the efficient control of Arthur A. Center. The first and third portions of the paper on the Tri-State region were presented by B. M. O'Harra, with R. H. Lowe substituting for co-author Illidge on the second part. Attention was drawn to the compact design attainable when "Sink and Float" was employed as a means of primary concentration. The teaching profession should welcome this concise correlation of data from such a renowned center of lead and zinc production. H. R. Hanley and C. R. Hay¬ward were much interested in the mechanics of charging, manufacture, and life of the retorts.

Jan 1, 1943

By A. M. E. Eilers

No one who has visited our Western mining districts, and studied the economical part of the beneficiation of the ores occurring all over that vast extent of country, can underrate the high importance of the above subject. By far the larger number of the districts which contain smelting ores, i.e., argentiferous and auriferous lead, or copper ores, are situated in the Great Basin, that great plateau between the Rocky Mountains on the east, and the Sierra Nevada on the west, almost the whole of which is comprised at present in the boundaries of Nevada, Utah, and part of Arizona. This region is essentially a barren country. The extreme dryness of the atmosphere permits but a scanty vegetation in the plains ; and even in the detached mountain-chains running through it-generally from N. to S., or from N. W. to S. E.-there are no trees found except dwarfed pines and mahogany, at the head of sheltered ravines, and a few cottonwoods and willows, which fringe the insignificant streams, before the water sinks in the arid plains. Nearly all the mountain-chains in this region are rich in silver ores. That class of these ores which is adapted to amalgamation, and rich in silver, has been worked with profit for more than ten years. But before the advent of the transcontinental railways, mining was restricted to these ores alone, and the consumption of fuel could be met with the scanty supply of forest trees in the immediate vicinity of the mining districts. Since, however, the Union and Central Pacific Railroads have brought the Great Basin nearer to the commercial centres of the East and the Pacific coast, thus reducing the expenses of freight and labor materially, other silver deposits, containing poorer ores in greater abundance, have been rapidly taken

Jan 1, 1873

By J. E. Owen

EXPERIMENTS on twenty-six drilling muds were carried out in an effort to establish a correlation between the apparent viscosity values for these muds as obtained with the Marsh funnel and the Stormer viscometer. This required a calibration of both instruments in which the density of the calibrating liquids was taken into account. Corresponding to the Marsh-funnel viscosity values that were obtained at 1500 c.c. in, 1 qt. out, equivalent Stormer viscosity values were obtained at a Stormer cylinder test speed of approximately 1100 r.p.m. Further, a linear relation between the logarithms of the Stormer cylinder test speed and the resulting apparent viscosity was found to exist for the muds tested for 200 to 1000 r.p.m. The significance of the slope of this linear relationship is discussed as it applies to the field use of drilling muds. It was concluded that Marsh-funnel viscosity measurements are expressible in absolute units that may be correlated with Stormer viscosity measurements, provided these are made in the manner described in the paper. INTRODUCTION The Marsh funnel is used almost universally to determine the viscosity of drilling muds in the field. The Stormer viscometer is commonly used to determine the viscosity of drilling muds in the laboratory. With the Marsh funnel, the time required for one quart of mud to flow out of the funnel originally filled with 1500 c.c. is often taken as the measurement of viscosity. With the Stormer instrument, drilling mud is tested at a cylinder speed of 600 r.p.m. and by means of calibration data the apparent viscosity is expressed in absolute units. Neither of these instruments is an absolute viscometer, and apparent viscosities obtained with either are influenced by the arbitrarily standardized conditions under which the measurements are made. It is disconcerting to the man in the field to be told that laboratory measurements of the viscosities of a group of muds differ widely, when he has already observed that they have identical viscosities when tested in the Marsh funnel. Drilling muds are not perfect fluids, therefore it is not surprising to find a lack of correlation between the apparent viscosity of a mud as determined with the Marsh funnel and that determined with the Stormer viscometer. The work described here was undertaken to develop a correlation between Marsh-funnel and Stormer viscosity measurements. To a considerable degree this has been accomplished. This study has resulted not only in a better understanding of the two viscosity instruments and the data obtainable from them, but has given considerable information regarding the viscosity characteristics of drilling muds. EXPERIMENTAL PROCEDURE Calibration of Harsh Funnel and Stormer Viscometer The first step in the experimental procedure was the calibration of the Marsh funnel and the Stormer viscometer with

Jan 1, 1941

By John H. Hollomon

The interpretation of notched-bar impact results has been a matter of controversy since the introduction of more or less standard tests by Fremont,' Charpy2 and others at the turn of the century. Many investigators3-l1 have contributed to the understanding of the significance of such tests. Several symposiums12,13 have been held with the express purpose of discussing the significance and interpretation of the impact tests. It is not the purpose of this paper to review all these contributions to the development of the knowledge on this subject, but rather to present an interpretation of notched-bar impact tests that appears to be in agreement with the available published data. Fundamentally, the interpretation discussed in this paper of the brittle failure of some steels in notched-bar impact test is similar to the qualitative analysis presented by Ludwik3 in the I92o's. Essentially, Ludwik stated that the function of the notch was to increase the tensile stress necessary for yielding, and that brittle failure occurred when the stress necessary for yielding exceeded the normal stress required for fracture (with no plastic deformation). Ludwik, however, had no means of measuring this normal stress necessary for failure of a plastically* undeformed metal, therefore he could not extend his interpretation beyond the qualitative stage. Kuntze14 and recently Mc-Adam15 attempted to determine the effect of deformation on the stress necessary for fracture by testing, in tension, specimens notched to various depths and angles. Because of the uncertainties in Kuntze's (and McAdam's) analysis of the stresses (which have been pointed out by Sachs16) and the inherent nonuniformity in the stress distribution in notched bars, the results of such investigations are difficult to interpret. A more direct method of measuring the effect of deformation on the fracture stress was suggested in previous papers"-18 and has been utilized in the present investigation. The present paper concerns itself primarily with the behavior of notched impact specimens. The effects of strain rate, temperature, and stress distribution, which are discussed with reference to the impact specimen, apply just as well to the behavior of metal at the bases of notches in any engineering structure. All that is necessary is to determine the stress distribution in the structure at the base of the notches.

Jan 1, 1944

By John H. Hollomon

The interpretation of notched-bar impact results has been a matter of controversy since the introduction of more or less standard tests by Fremont,' Charpy2 and others at the turn of the century. Many investigators3-l1 have contributed to the understanding of the significance of such tests. Several symposiums12,13 have been held with the express purpose of discussing the significance and interpretation of the impact tests. It is not the purpose of this paper to review all these contributions to the development of the knowledge on this subject, but rather to present an interpretation of notched-bar impact tests that appears to be in agreement with the available published data. Fundamentally, the interpretation discussed in this paper of the brittle failure of some steels in notched-bar impact test is similar to the qualitative analysis presented by Ludwik3 in the I92o's. Essentially, Ludwik stated that the function of the notch was to increase the tensile stress necessary for yielding, and that brittle failure occurred when the stress necessary for yielding exceeded the normal stress required for fracture (with no plastic deformation). Ludwik, however, had no means of measuring this normal stress necessary for failure of a plastically* undeformed metal, therefore he could not extend his interpretation beyond the qualitative stage. Kuntze14 and recently Mc-Adam15 attempted to determine the effect of deformation on the stress necessary for fracture by testing, in tension, specimens notched to various depths and angles. Because of the uncertainties in Kuntze's (and McAdam's) analysis of the stresses (which have been pointed out by Sachs16) and the inherent nonuniformity in the stress distribution in notched bars, the results of such investigations are difficult to interpret. A more direct method of measuring the effect of deformation on the fracture stress was suggested in previous papers"-18 and has been utilized in the present investigation. The present paper concerns itself primarily with the behavior of notched impact specimens. The effects of strain rate, temperature, and stress distribution, which are discussed with reference to the impact specimen, apply just as well to the behavior of metal at the bases of notches in any engineering structure. All that is necessary is to determine the stress distribution in the structure at the base of the notches.

Jan 1, 1944

By C. B. E. Douglas

In an interesting article on evaluation of the water problem at Eureka, Nev., W. T. Stuart describes how a difficult water problem, or one phase of it, may be evaluated by means of a small scale test. Test data are plotted by a method rendering, under certain conditions, a straight-line graph that can be projected to show how much the water table will be lowered by pumping at any specified rate for a given time. A formula is then used to determine the size of opening, or extent of workings, necessary to provide sufficient inflow to enable pumping to be maintained at that rate. At first glance this might seem the answer to a miner's prayer, but a word of caution is in order. It may not be the whole answer. Moreover, results obtained by the method described are reliable only for conditions approximating those assumed. Even where conditions do not meet this requirement, however, it may be possible to draw helpful inferences from the results, perhaps enough to facilitate another approach to evaluation of a problem.

Nov 1, 1955

By J. K. Brooke, R. M. Dreyer

For mining engineers and geologists accustomed evaluating metallic mineral deposits, the problems inherent in determining the worth of non- metallic industrial minerals deposits for the chemical processing industry often comes as a rude shock. These difficulties lead to different, and often much greater, problems in financing the development of such deposits than would normally be anticipated for economic sources of copper, molybdenum, iron, nickel and other metallic minerals. The chemical processing industry is one of the two principal consumers of non-metallic mineral production, the other being the construction industry. In the last decade, the volume of industrial minerals consumed by chemical processers has in- creased by 30%, and further increases are forecasted. To satisfy this growing demand, many new sources of these minerals must be found and developed. For those of us who will become involved in such work, the following brief analysis of the major problems to be expected may be of help in developing a realistic approach to evaluating and financing such a deposit and selling the product to consumers.

Jan 8, 1962

By R. V. Wheeler, F. V. Tideswell

Coal being essentially a complex conglomerate of plant remains that have undergone decay and interaction in varying degree, it is understandable that attack on the problem of its chemical constitution by normal chemical methods should have been attended by but slight success. The obvious method of fractionation by means of selective solvents has yielded a few results of some value. Thus it is known that from many, if not all, bituminous coals organic solvents extract small percentages of oxygenated compounds, which are generally assumed to have been derived from the original plant resins, though a more probable source would appear to be the glycerides and waxes of the plants. Organic solvents also extract small proportions of hydrocarbons, the origin of which is doubtfill. Of more importance is the action of such pseudo solvents as pyridine, which, in addition to their specific solvent action, have the power of resolving and simplifying a portion of the coal substance in a manner analogous to that by which water can disperse a hydrogel. Early work with pyridine indicated that coals are composed of two main classes of compounds: the one soluble in organic solvents, the other capable of being partly dispersed in pyridine. Recent work has shown that the dispersion is colloidal in character and that it may be accomplished by the use of solvents less active than pyridine, such as benzene or chloroform, provided that the temperature of extraction is raised or the coal has been suitably preheated. The action of reagents on coal has proved of little value in the past, but some insight into the constitution of coal has been given by regulated destructive distillation. Thus, it has been shown that coals behave on distillation as though composed of two main classes of compounds: the one readily decomposable at moderate temperatures (yielding mainly the paraffins), the other undergoing a preliminary early decomposition (with evolution of oxides of carbon and water) and yielding a residue that at higher temperatures breaks down, yielding much hydrogen.

Jan 1, 1925

Open pit operators have found the dropball crane a great aid in cutting costs and hazards of secondary breakage. But a common problem in using the unit has been with the shock absorbing connection from the cable to the dropball itself. Blair Limestone Div. of Jones & Laughlin Steel Corp. has come up with a new connector to solve this problem and to save the time usually lost in frequent replacement of the cable slings, chains, or old tires hitherto used at this point. J&L personnel solved the problem by combining three ideas from other dropball crane owners. Their unit consists of two long U-bolts running through a heavy spring originally made for use on a vibrating screen. Heavy duty swivels attached to the bent ends of the U-bolts provide attachment points for the crane cable and the special chain to the dropball. Connection at the dropball is by an eyelet welded directly into the mass of steel, and a special alloy chain is connected to this with a clevis.

Nov 1, 1955

By Harvey T. Kennedy, Erasmo T. Guerrero

In this investigation, the effects of the surface tension of brine and the interfacial tension between oil and brine on the recovery of oil by water flooding of Woodbine sand were studied. Variation of the surface forces was obtained by the addition of small amounts of commercially available chemicals, which did not appreciably change the surface tension of the oil. Fourteen surface-active agents, which were found to be effective in lowering the surface and interfacial tensions, were further investigated to determine the extent of their adsorption by the Woodbine Sand cores. It was found that adsorption was sufficient in all cases to prevent any substantial concentration of chemical at the flood front. The flow system was, therefore, brought to equilibrium with definite concentrations in the flood water by circulation before each test. The data thus have no relation to the results obtainable by using the chemicals in full-scale water floods, but should be helpful in predicting the performance of surface-active chemicals which are not lost by adsorption, if such can be found. Flow tests were made on various concentrations 01 six chemicals selected from the 14 investigated for surface activity and adsorption, with and without the presence of gas in the core. It was found that lowering the interfacial tension had a tendency toward reducing the recovery of oil and that the reduction of the surface tension of the brine had little or no effect. A substantial beneficial effect, however, resulted from the presence of gas in the cores. INTRODUCTION The movement and distribution of crude oil in a porous medium are governed by such factors as surfacc and interfacial tensions, viscosity of the oil, pore size configuration, and wetting characteristics of the medium. The knowledge of the part played by each of these factors is limited. If the oil is assumed to be distributed in droplets whose movement is restricted only by the pore openings, the following equation may be applied: ?P= 2r(1/r1-1/r2) .......(1) In this equation ?P is the pressure drop necessary to force an oil droplet through the construction, r is the oil-brine interfacial tension, r1 is the radius of curvature at the forward and smaller end of the deformed droplet, and r2 is the larger radius. This equation neglects the effects of viscosity, as a static or near static condition is assumed. It would apply equally well for a bubble of gas in liquid if r were taken as the surface tension of the liquid. This relation shows that droplets can be more easily displaced if the interfacial tension is decreased. The displacement of oil from the surface of a sand particle may also be considered from the basis of selective wetting assuming that the oil droplets are smaller than the pore constructions and that the sand surfaces may be considered as planes. A relation of the forces involved can be derived from Dupre's equation to give: Wos = row (1 +Cos?)........(2) where Wos = Work of adhesion between the sand and the oil row = Oil-water interfacial tension 0 = Contact angle ros = Oil-solid interfacial tension This equation would indicate that the displacement of oil from a sand surface is more readily obtained when the water-oil interfacial tension is reduced provided the contact angle remains constant. Evidently. however, the values of Cos 0 may vary along with the water-oil interfacial tension due to changes in the other surfaces of the system. The above theoretical considerations suggest that perhaps even the static behavior of a dispersed oil phase in a porous medium is considerably more complex than may be inferred by the simple systems

Jan 1, 1955

By C. E. Gregor

Current modern trends in mining show that hoisting shafts are being expected to fulfill an important ventilation function. However, where rigid guides and supporting structures are mandatory, ventilating power costs will be high, unless the transverse structures (shaft sets) are streamlined. In the complex configurations of hoisting shaft furniture patterns, either for original design purposes, or for the adjustment of an existing shaft, a useful degree of aerodynamic optimization, by reduction of overall resistance, can only be effectively evaluated by model studies in a special purpose wind tunnel, as described herein. The author has shown that, in one particular case, an annual power cost saving of more than $50,000 was indicated, mainly by the selection of buntons and dividers of improved aerodynamic profile. Mining companies are now realizing the advantages that can accrue through proper attention to the aerodynamic aspects of mine shaft design. With progressive demands of civilization for an ever greater supply of coal and metals, and with greater depths of working, higher rates of extraction, advanced degrees of mechanization, and improved standards of environmental health and work comfort, shaft design is becoming an extremely important aspect of mining operations. In addition, because the cost of sinking, equipping, maintaining, and operating a mine shaft is continually increasing, there is a greater incentive to use a shaft for the maximum duty, consistent with economy. The current natural tendency is therefore to sink multi-purpose shafts, to install as much equipment as is safely and effectively possible to operate; and then, over and above all these functions, to use the shaft to its maximum possible ventilating capacity, as a major, or auxiliary, airway. Clearly, this latter additional function can only be employed at a significant cost, since the most economic operating airways are smooth-lined and devoid of internal rigid furniture. It is true that hoisting functions can be carried out in such shafts if rope guides are used, but only under restricted conditions. Where a manway must be provided, fixed guides and supporting structures are mandatory, and ventilation costs become accentuated. It is generally recognized that the prime element of cost in ventilating a mine can be represented by the power required to drive the fan unit. It is also known that the largest component of total mine resistance is that of the downcast shaft, where this shaft is equipped with rigid guides for hoisting purposes. The pressure loss varies directly as the resistance of the shaft, in accordance with a modification of the Atkinson equation p = RQ2/3600. Furthermore, the air horsepower drawn is given by the equation AHP = (p x 0/33000. Therefore, the power required to drive the fan unit varies as the resistance of the shaft and the cube of the air quantity. Clearly, if Q is a specified minimum, then the only way to minimize the pressure loss (and the power costs) is to reduce the shaft resistance R. (See Table I for a list of symbols.) The resistance to airflow of a shaft equipped with rigid guides and supporting structural sets consists of two components which, for the sake of analytical simplicity, can be regarded as additive. 1) Frictional resistance, represented by the viscous frictional forces within the boundary layer, as air flows along the surfaces of the walls and the periphery of the vertical elements (guides, pipes, cables, ropes, etc.) of the furniture. The friction loss can be readily calculated by the Atkinson equation, after selection of a suitable value for the friction factor K from the McElroy compilation.' KxLxO xQ2 5.2 xA3 Friction loss can be minimized by lining the shaft walls (particularly if of circular or elliptical cross-section) with a smooth continuous lining of monolithic concrete.

Jan 1, 1968

By Warren E. Wilson

The transportation of solids in pipe lines is a matter of deep concern in many fields of engineering. Much experimental and theoretical work has been done in an effort to devise means of designing pipe lines that will satisfactorily transport given solid loads, but there remain many unanswered questions in this field. The present discussion will be confined to a rather narrow portion of this field and will endeavor to point out certain factors that may well be considered when comparing the relative economy of various installations. The data have been presented previously but a new interpretation was prompted by a suggestion in a paper by Professor O'Brien,1 of the University of California, that the friction work per pound of dry material should be considered in determining the relative efficiency of various installations. This method of considering relative economy in various pipe lines leads to some rather unexpected conclusions, especially with reference to the economy of large pipe sizes. A previous suggestion by the writer3 concerning the economy of a rather peculiar type of installation is strengthened. It is not believed that the considerations set forth herein apply directly in the case of dredge lines, where economy is largely a matter of a high rate of pumping in order to reduce the unit overhead charge. However, it is believed that some value may be gained from these considerations in cases involving the disposal of a fixed amount of solids per unit time with a minimum expenditure of energy and at a low cost for replacement of worn pipes. Such situations are not uncommon in the disposal of mill tailings and the disposal of solids deposited from irrigation waters. It was found that data leading to general conclusions for various pipe sizes were not available but certain definite trends are indicated and methods of analyzing available data are presented in the hope that they may be of assistance in preparing engineering estimates on designs of pipe lines of the type described above. Energy Requirements Carrying out the idea contained in ' Professor O'Brien's paper concerning the method of computing energy required to transport solids, the following analysis leads to some rather interesting conclusions. Let us represent: Solids load in term of mass per unit time by W Ratio of mass of suspended solids to total mass of mixture flowing by p Volume rate of flow of mixture by Q Length of pipe by L slope of energy grade line by s Mass per unit volume of solids by m. Mass Per unit volume of liquid by mi Mass per unit volume of mixture by mm We then have as an expression for the solids load: W = pQmm The work done per unit mass, P, on the mixture to move it along the pipe the

Jan 1, 1948

By Warren E. Wilson

The transportation of solids in pipe lines is a matter of deep concern in many fields of engineering. Much experimental and theoretical work has been done in an effort to devise means of designing pipe lines that will satisfactorily transport given solid loads, but there remain many unanswered questions in this field. The present discussion will be confined to a rather narrow portion of this field and will endeavor to point out certain factors that may well be considered when comparing the relative economy of various installations. The data have been presented previously but a new interpretation was prompted by a suggestion in a paper by Professor O'Brien,1 of the University of California, that the friction work per pound of dry material should be considered in determining the relative efficiency of various installations. This method of considering relative economy in various pipe lines leads to some rather unexpected conclusions, especially with reference to the economy of large pipe sizes. A previous suggestion by the writer3 concerning the economy of a rather peculiar type of installation is strengthened. It is not believed that the considerations set forth herein apply directly in the case of dredge lines, where economy is largely a matter of a high rate of pumping in order to reduce the unit overhead charge. However, it is believed that some value may be gained from these considerations in cases involving the disposal of a fixed amount of solids per unit time with a minimum expenditure of energy and at a low cost for replacement of worn pipes. Such situations are not uncommon in the disposal of mill tailings and the disposal of solids deposited from irrigation waters. It was found that data leading to general conclusions for various pipe sizes were not available but certain definite trends are indicated and methods of analyzing available data are presented in the hope that they may be of assistance in preparing engineering estimates on designs of pipe lines of the type described above. Energy Requirements Carrying out the idea contained in ' Professor O'Brien's paper concerning the method of computing energy required to transport solids, the following analysis leads to some rather interesting conclusions. Let us represent: Solids load in term of mass per unit time by W Ratio of mass of suspended solids to total mass of mixture flowing by p Volume rate of flow of mixture by Q Length of pipe by L slope of energy grade line by s Mass per unit volume of solids by m. Mass Per unit volume of liquid by mi Mass per unit volume of mixture by mm We then have as an expression for the solids load: W = pQmm The work done per unit mass, P, on the mixture to move it along the pipe the

Jan 1, 1948

By Fred E. Wright

In a recent paper1 a detailed discussion is given of the possibilities of using polarized light in the examination of opaque substances. The factors underlying the problem are there treated from the viewpoint of the electromagnetic theory of light and the presentation is necessarily mathematical in form. The practical methods resulting from the discussion are, however, relatively simple and promise to be of value in the determination of certain ores and metals when studied in reflected light under the microscope. The results of the theoretical investigation referred to may be summarized by stating that in general the optical constants, such as refractive indices and absorption indices, cannot be satisfactorily ascertained on small, random, polished sections; that the use of polarized light enables the observer ordinarily to determine whether the crystal plate is iso-tropic or anisotropic, and also to ascertain the degree of anisotropism; that for this purpose methods based either on the contrast in intensity between the two reflected components of non-polarized light, or on the amount of rotation of the plane of polarization, on reflection, of vertically incident, plane-polarized light, may be employed; that methods based on the phase difference between the two reflected components are in general of little value because of the small differences in phase which ordinarily result for a relatively large change in birefringence or biabsorption. The measurement of the optical properties of transparent minerals, even in minute irregular grains, is a simple task with modern petrographic microscope methods and is accomplished by petrologists as part of their ordinary routine work. But the determination of the optical constants of opaque substances is difficult and is rarely attempted; all observations are necessarily made in reflected light and are restricted commonly to the determination of color, of hardness, of the character of crystallization, and of the behavior of the mineral or metal plate toward reagents and abrasives. With practice, the observer becomes expert in recognizing details of this kind and determines with a high degree of certainty the different

Jan 1, 1920

By Niles E. Grosvenor

Complex underground mining problems are increasing as mining depths increase. Many of these problems have been solved mainly by unsystematized trial-and-error methods based on individual experiences. This system of solving mining problems has resulted in over-cautiousness whereby a large quantity of our mineral resources are left in the ground, or not enough caution with a resultant failure of the underground opening. From studies made of physical properties of rocks, large volumes of data have been collected as a source of finding correlation among the different physical tests or among certain rock types. Most of these studies have not produced satisfactory results because the investigator usually arrives at the conclusion that the results are too scattered to be reliable. The reasons usually given as the cause of this scatter are that rocks are nonhomogenous and non-isotropic. Similar types of poor correlation that were being obtained from compression tests in the Colorado School of Mines Mining Research Laboratory led to an investigation of a good method of making a simple unidirectional compressive strength test.

Jan 1, 1963

By AIME

Lewis E. Young, who will formally assume his duties as President of the AIME at the Annual Meeting in San Francisco in February 1949, was born in Topeka, Kansas, Oct. 1, 1878. Dr. Young received his engineering education at the Pennsylvania State College, where he received the degree of B.S. in Mining Engineering in 1900. In 1904 he received his E.M. degree from Iowa State College, and in 1915 a Ph.D. from the University of Illinois. In 1947, the University of Missouri conferred on Dr. Young honoris causa the degree of Doctor of Engineering. His active career began in the educational field as instructor at Iowa State College, and following that, as professor of mining at the Colorado School of Mines, 1903-1907. While he was at Golden he launched hi: career as consulting mining engineer for various Western companies. Leaving the Colorado School of Mines, Dr. Young returned to Missouri, serving the University of Missouri as director of the School of Mines and Metallurgy at Rolla, 1907-1914. Leaving Missouri he went to the University of Illinois as assistant professor of business organization, 1915-1918, during which period he wrote exhaustive treatises on mine subsidence and mine taxation, which won for him the degree of Ph.D. before mentioned.

Jan 1, 1949

By J. L. Huitt, B. B. McGlothlin, D. K. Lowe

Published correlations for the principal aspects of hydraulic fracturing were combined into a digital computer program to facilitate the study of interrelated variables. The computer program includes individual relationships for fracture width during pumping, fracture area generated, propping agent embedment, flow capacities of propped fractures and transport of propping agents in horizontal fractures. The effects of more than 20 treatment and formation parameters on the predicted results of hydraulic fracturing treatments were studied. The effects of these parameters were determined for (I) fracture width during injection, (2) fracture width after the overburden comes to rest on the propping agents, assumed not to be crushed, (3) generated and propped fracture area, (4) location and concentration of propping agents in the fracture when injection ceases, (5) flow capacities of the various propped sections of the fracture and (6) expected increase in the well productivity. The effects of propping agent, formation and fracturing fluid parameters on well productivity are discussed. The parameters that were found to have the most pronounced effects on hydraulic fracturing treat~nents are injection rate, treatment volume, fracturing fluid coefficient, size and amount of propping agent, spearhead volume, well drainage radius and formation capacity. INTRODUCTION Many correlations have been published for predicting effects of various parameters that are considered in the design of hydraulic fracturing treatments. The Carter equation' can be used to predict generated fracture radius as a function of fracture width, fracturing fluid leakoff and other parameters. Fracture width can be determined by use of the Perkins and Kern correlation' in which the fracture width is related to the fracture radius, fluid injection rate and certain formation and fracturing fluid parameters. Wahl and Lowe et aL4 have reported methods of predicting the location of propping agents in fractures when pumping ceases. The former study is applicable to the case where the ratio of propping agent diameter to fracture width is less than 0.1. The latter is applicable when this ratio is greater than 0.1. These studies showed that the propping agent placement in horizontal-radial fractures depends principally on how the individual particles are transported in the fracture by the carrying fluid. Particle transport in fractures is determined by local fluid velocity in the fracture, fluid and particle properties, and the size of the particle relative to the fracture width. The distribution of propping agents, effective overburden pressure and formation rock strength control the propped fracture width6 by controlling the extent to which the propping agent particles embed into the fracture faces. From the distribution of propping agents and the propped fracture width, fracture flow capacities can be calculated or the various regions of the fracture. The flow capacities and the radial extent of these regions can be combined with reservoir information to predict the productivity increases for fractured wells. In all these studies, the effects of certain treatment and/ or reservoir parameters on one facet of fracturing can be predicted only if other facets which the parameters affect are fixed. For instance, fracture width and radius are interrelated; that is, to calculate the value of one, the value of the other must be known. Also, some parameters influence more than one aspect of fracturing. For example, prop-pant transport is a function of both fracture width and fluid viscosity. but fracture width is itself a function of fluid viscosity. Since these calculations are complex and the parameters interrelated, it is not possible to write an equation with which the over-all effects of treatment parameters can be solved explicitly. For these reasons, the correlations for determining the effects of the parameters which are most significant in hydraulic fracturing treatments have been incorporated into a digital computer program. COMPUTER PROGRAM The program, which was written for an IBM 7094 computer, can be used to predict results of most of the combinations and values for the treatment parameters that are ordinarily considered for fracturing treatments. A spearhead of fracturing fluid and a propping agent-carrying fluid with different fluid properties can be taken into account. Also, the total volumes and relative amounts of the spearhead and carrying fluids can be varied. Two different propping agents (as used in tail-in operations) and a wide range of formation properties and injection rates are considered. The computer program (Fig. 12) consists of several sets of calculations. First, the final flooded fracture radius and average fracture width at the cessation of pumping are calculated. This is done by simultaneous solution of the Perkins and Kern fracture width equation and the Carter equation for flooded fracture radius (equations used in the computer program appear in the Appendix). The next step is to determine local fluid velocity in the fracture as a function of time and radius. Since it is not possible to write this function in closed form expression, a table of velocity values is generated by the program and stored for subsequent use. The time span from the beginning of