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By Ellsworth Daggett

THE system of pitwork used with the Cornish pumping engine, and which, for want of a better name, we may call the Cornish system of pitwork, consists essentially of a series of plunger-pumps, situated one below the other, in a shaft or incline, the plungers being operated by a rod called the pitman or pump-rod, which must, therefore, extend from the motor at the surface to the lowest pump. The subject will be best understood by reference to Fig. 1, in which k may be taken as typical of the ordinary plunger-pump, in which S is the suction pipe, u, the lower valve situated in the H piece T; v, the upper valve situated in the valve-box O; I, the mounting-pipe or column through which the water is carried to the surface; L, the plunger-barrel, and P, the plunger working through the stuffing-box, and actuated by the branch rod r, which is attached to the main pump-rod, R. The action of the pump is as follows : On the ascent of the plunger, acting now by suction, the water flows from the reservoir through the valve u, into the barrel, L. On the descent of the plunger, the water is forced up through the valve, v, which acts merely as a check valve, into the pipe, I, through which it is, at each successive down-stroke of the plunger, forced upward until it reaches the next reservoir, from which it is taken by another similar pump, and so on until it reaches the surface or drain-tunnel. The pump is obviously only single-acting, i. e., it acts only during one-half of its entire stroke. The weight of water raised during the up-stroke of the plunger is equal to the weight of a column of water whose base is the area of the plunger and whose height is the mean distance between the plunger-face and the surface of water in the lower reservoir, a distance not usually exceeding 12 or 15 feet, and which obviously cannot under the most favorable circumstances, exceed 25 or 30 feet. The water forced up during the down-stroke is equal to the weight of a column of water whose base equals the area of the plunger and whose height equals the length of the mounting column above the mean position of the plunger-face, say from 150 to 300 feet. Almost all of the useful work, therefore, is done on the de-

Jan 1, 1879

By Paul B. Crawford, Melvin B. Burton

A procedure is given which describes how the gelatin model may be-used to study the effect of mobility ratio on areas swept before and after breakthrough in fluid injection programs. Differences in mobility ratio are achieved through the proper selection of electrolytes for the invading and displaced fluids. The progress of the front is followed visually and photographed or traced onto plates (it desirrd positions before and after breakthrough. Example applications to standard flooding pattrrrls, vertically and horizontally fractured reservoirs and the in-situ combustion recovery process ore presentrtl. INTRODUCTION In water flooding, gas cycling and other fluid injection programs, the sweep efficiencies and swept areas after breakthrough are known to depend on the mobility ratio. Previous investigators have described the application of the X-ray technique, stepwise use of the poten-tiometric model and numerical procedures for studying these problems.1,3,4,6 Others have described the application of the electrolytic and gelatin models for studying sweep efficiencies at a mobility ratio near one.2,5,7,8,9 It is a purpose of this paper to describe a procedure for using the gelatin model to study swept areas at mobility ratios other than one. PROCEDURE The technique developed for gelatin model studies is composed of four phases: the construction of the model, the selection and preparation of electrolytes, the running procedure, and the recording of the swept areas. Each phase is discussed independently in order that the over-all procedure will not be confused. Construction of the Model Since some of the advantages of the gelatin model are its economy, and speed, it is desirable to have as small a model as is possihle without creating error through the inability to follow visually the progress of the colored front. For this work, it was found that a distance of 4 to 6 cm between salt bridges is convenient to use. Since the size is relatively small, it follows that the gelatin layer must be thin. For this study, 1/8 in. was used as the gelatin thickness in each case. The model was constructed on a flat 1/2-in. plastic (Lucite) base with boundaries constructed from 1/8-in. plastic strips. The well pattern to be studied was outlined with the 1/8-in. plastic strips on the base with narrow apertures (- - 2 mm) in the boundary to represent the wells. Behind each well opening, a rectangular area of some 8 to 12 sq cm was outlined by additional plastic strips. The plastic strips were fastened to the base by any suitable non-conducting cement so that the boundary system was liquid tight. The electrodes were constructed from sheet copper in a fashion that would allow easy removal. In this work, it was found that by simply forming a large U-shaped clip from a strip of the copper that would slip over the boundary was satisfactory and easily removed. The electrodes should be as wide as one of the sides of the rectangular area behind the well opening. If a fractured pattern is studied, fractures inside the well pattern may be represented by strips of copper of the desired shape mounted in some easily removable fashion. The reason that apertures (or salt bridges) were used for wells instead of metallic electrodes is that the plating of the electroactive ion at the cathode and the introduction of an electroactive ion at the anode has a detrimental effect on the swept area outline if these operations \rere permitted to be carried on within the well pattern itself. The anode may be used within the well pattern

Jan 1, 1957

By Irwin Politziner F. M. Townsend, L. S. Reid

Recently published data indicate that the water vapor content of a gas, as determined by dew point measurement, is inaccurate when the gas has been dehydrated with diethylene glycol. Water vapor contents and dew point measurements of gases dehydrated with triethylene glycol have been obtained in this investigation in order to determine the magnitude of a similar error, if one exists. Experimental data show a very low concentration of triethylene glycol vapor in gases dehydrated at atmospheric ternperatures and pressures ranging from 500 to 2,500 psia and that the accuracy of dew point measurements is not impaired by the presence of triethylene glycol vapor. INTRODUCTION Nearly all natural gas transported by pipe line to northern and eastern markets must be dehydrated to a low residual moisture content in order to assure maximum transmission efficiency and continuous delivery under peak load conditions. Without dehydration, water vapor often condenses in pipe lines in quantities sufficient to restrict the flow3 and. under low atmospheric temperature conditions, gas hydrates may form and plug gathering. transmission and distribution facilities. Gas may be dehydrated to pipe line specifications by a number of methods. The most widely used methods are (1) adsorption of water vapor on a solid dessicant such as activated bauxite, alumina or silica gel,' (2) by absorption of water vapor by either diethylene or triethylene glycol-water solutions of high glycol concentration,11,13 and (3) by simultaneous expansion-refrigeration of very high pressure gas in which gas hydrates are purposely formed and then quickly decomposed.'"" Of these three fundamental methods, the adsorption and absorption processes provide the bulk of the dehydrated gas moving to market at the present time. The expansion-refrigeration process. a relatively new development, is gaining in favor because of its simplicity but is sharply limited in its application by available differentials between source and gathering system pressures. Where gas sale or purchase contracts contain a maximum water vapor content specification, the water vapor content is usually determined by measuring the dew point of the gas at system pressure. Knowing he dew point temperature and the system pressure, the water vapor content of the gas is determined from a graphical correlation of saturation, or dew point, temperature us the water vapor content of the gas which is customarily expressed as pounds of water vapor per million cubic feet of gas measured at 14.7 psia and 60°F. Dew points are usually measured with the U. S. Bureau of Mines Dew Point Tester. shown in Fig. 1, which is well suited for use at temperatures lower than 120°F and at pressurei; ranging from atmospheric to 3,000 psi. A number of correlations of dew point temperatures vs water vapor content are available,3,5,15 the most recent of which is that by McCarthy et al.4 Where natural gas is dehydrated by absorhing contact with concentrated diethylene glycol-water solutions. considerable difficulty is experienced in observing water dew points due to condensation of liquid hydrocarbon and glycol films on the mirrored surface of the dew point tester. Riesenfeld and Frazierl2 report that the dew point method for determining the water vapor content of a dietllylene glycol-treated natural gas is in error due to the glycol content of the condensate formed on the tester mirror, and that the water vapor content of the gas tested is actually lower than that indicated by the standard' dew point-water vapor content correlation. Extensive experience in testing gases dehydrated by concentrated triethylene glycol-water solutions lias not revesled difficulties of the nature cited above for diethylene glycol. Water dew points are clear and sharp and condensation of

Jan 1, 1951

By L. S. Reid, Irwin Politziner F. M. Townsend

Recently published data indicate that the water vapor content of a gas, as determined by dew point measurement, is inaccurate when the gas has been dehydrated with diethylene glycol. Water vapor contents and dew point measurements of gases dehydrated with triethylene glycol have been obtained in this investigation in order to determine the magnitude of a similar error, if one exists. Experimental data show a very low concentration of triethylene glycol vapor in gases dehydrated at atmospheric ternperatures and pressures ranging from 500 to 2,500 psia and that the accuracy of dew point measurements is not impaired by the presence of triethylene glycol vapor. INTRODUCTION Nearly all natural gas transported by pipe line to northern and eastern markets must be dehydrated to a low residual moisture content in order to assure maximum transmission efficiency and continuous delivery under peak load conditions. Without dehydration, water vapor often condenses in pipe lines in quantities sufficient to restrict the flow3 and. under low atmospheric temperature conditions, gas hydrates may form and plug gathering. transmission and distribution facilities. Gas may be dehydrated to pipe line specifications by a number of methods. The most widely used methods are (1) adsorption of water vapor on a solid dessicant such as activated bauxite, alumina or silica gel,' (2) by absorption of water vapor by either diethylene or triethylene glycol-water solutions of high glycol concentration,11,13 and (3) by simultaneous expansion-refrigeration of very high pressure gas in which gas hydrates are purposely formed and then quickly decomposed.'"" Of these three fundamental methods, the adsorption and absorption processes provide the bulk of the dehydrated gas moving to market at the present time. The expansion-refrigeration process. a relatively new development, is gaining in favor because of its simplicity but is sharply limited in its application by available differentials between source and gathering system pressures. Where gas sale or purchase contracts contain a maximum water vapor content specification, the water vapor content is usually determined by measuring the dew point of the gas at system pressure. Knowing he dew point temperature and the system pressure, the water vapor content of the gas is determined from a graphical correlation of saturation, or dew point, temperature us the water vapor content of the gas which is customarily expressed as pounds of water vapor per million cubic feet of gas measured at 14.7 psia and 60°F. Dew points are usually measured with the U. S. Bureau of Mines Dew Point Tester. shown in Fig. 1, which is well suited for use at temperatures lower than 120°F and at pressurei; ranging from atmospheric to 3,000 psi. A number of correlations of dew point temperatures vs water vapor content are available,3,5,15 the most recent of which is that by McCarthy et al.4 Where natural gas is dehydrated by absorhing contact with concentrated diethylene glycol-water solutions. considerable difficulty is experienced in observing water dew points due to condensation of liquid hydrocarbon and glycol films on the mirrored surface of the dew point tester. Riesenfeld and Frazierl2 report that the dew point method for determining the water vapor content of a dietllylene glycol-treated natural gas is in error due to the glycol content of the condensate formed on the tester mirror, and that the water vapor content of the gas tested is actually lower than that indicated by the standard' dew point-water vapor content correlation. Extensive experience in testing gases dehydrated by concentrated triethylene glycol-water solutions lias not revesled difficulties of the nature cited above for diethylene glycol. Water dew points are clear and sharp and condensation of

Jan 1, 1951

By Fred H. Poettmann

An equation has been derived for use in calculating the sandface pressure of flowing gas wells in which the variation of the compressibility factor of the gas with pressure is taken into consideration. This variation due to compressibility has been put into both graphical and tabular form. Comparison of calculated results with field measured results were made on 20 dry gas wells from a given field and 11 distillate wells from different fields. The agreement between calculated and observed results is good. In addition to their use in the calculation of sandface pressures of flowing wells, the factors can be used in the direct calculation of the static bottom hole pressure of gas wells, the capacity of gas transmission lines, and in the calculation of the theoretical isothermal horsepower necessary to compress a natural gas. Examples demonstrating the use of various equations are given. INTRODUCTION Many of the equations used to calculate the sandface pressures of flowing gas wells from well head data do not take into consideration rigorously the deviation of the natural gas from ideal gas behavior. For low pressure wells this error is not serious. For high pressure wells flowing at high rates this error call be serious. Sandface pressures are used to determine the producing capacity of gas wells as shown in the U. S. Bureau of Mines' Monograph 7.' An equation has been derived for use in calculating the sandface pressure of flowing gas wells in which the variation of compressibility of the gas with pressure is taken into consideration. This variation due to compressibility has been put into both graphical and tabular form. In addition to their use in the calculation of sandface pressures, these "factors" can also be used to give a direct solution in the calculation of the static bottom hole pressure of gas wells, the calculation of the capacity of gas transmission lines, and the calculation of the isothermal horsepower necessary to compress a gas. CALCULATION OF SANDFACE PRESSURE OF FLOWING GAS WELLS The starting point in the derivation of any specific flow equation is an energy balance on the fluid flowing between any two points in the system. This energy balance based on one pound of flowing fluid is expressed by the well known general flow equation: jp~- vdP+ Ah + -— + wt + r. = o . . (i) P, 2 g. where V = specific volume of flowing fluid in cu ft/lb P = pressure in Ib per sq ft ah = difference in height above datum plane in, ft ------ = kinetic energy change of one lb of flowing fluid, v equals velocity in ft/sec and g. = 32.174 Ws = Work done by fluid while in flow (similar to shaft work in driving a turbine) W, = Znergy losses due to the irreversibilities of the flowing fluid This equation contains no limiting assumptions and can be made the basis of any fluid flow relationship by limitations or substitutions. The energy units are usually expressed in term* of foot-pounds of energy per pound of mass. Considering the vertical flow of gas through tubing or annu-lus, the external work W, done by the gas is zero. Let us also assume that the value of the kinetic energy function is small and can be neglected. Equation (1) is reduced to: /P: VdP + Ah+Wt = 0 .......(2) Pi The energy loszes PI can be expressed in terms of the well known Fanning equatinn: 4/Lf: rf = —— ............(3) 2g,D where L = the integrated average velocity in ft/sec of vertical flow in a uniform flow string -~h = L v = the integrated average velocity in ft/sec / = dimensionless correlating function

Jan 1, 1951

By Fred H. Poettmann

An equation has been derived for use in calculating the sandface pressure of flowing gas wells in which the variation of the compressibility factor of the gas with pressure is taken into consideration. This variation due to compressibility has been put into both graphical and tabular form. Comparison of calculated results with field measured results were made on 20 dry gas wells from a given field and 11 distillate wells from different fields. The agreement between calculated and observed results is good. In addition to their use in the calculation of sandface pressures of flowing wells, the factors can be used in the direct calculation of the static bottom hole pressure of gas wells, the capacity of gas transmission lines, and in the calculation of the theoretical isothermal horsepower necessary to compress a natural gas. Examples demonstrating the use of various equations are given. INTRODUCTION Many of the equations used to calculate the sandface pressures of flowing gas wells from well head data do not take into consideration rigorously the deviation of the natural gas from ideal gas behavior. For low pressure wells this error is not serious. For high pressure wells flowing at high rates this error call be serious. Sandface pressures are used to determine the producing capacity of gas wells as shown in the U. S. Bureau of Mines' Monograph 7.' An equation has been derived for use in calculating the sandface pressure of flowing gas wells in which the variation of compressibility of the gas with pressure is taken into consideration. This variation due to compressibility has been put into both graphical and tabular form. In addition to their use in the calculation of sandface pressures, these "factors" can also be used to give a direct solution in the calculation of the static bottom hole pressure of gas wells, the calculation of the capacity of gas transmission lines, and the calculation of the isothermal horsepower necessary to compress a gas. CALCULATION OF SANDFACE PRESSURE OF FLOWING GAS WELLS The starting point in the derivation of any specific flow equation is an energy balance on the fluid flowing between any two points in the system. This energy balance based on one pound of flowing fluid is expressed by the well known general flow equation: jp~- vdP+ Ah + -— + wt + r. = o . . (i) P, 2 g. where V = specific volume of flowing fluid in cu ft/lb P = pressure in Ib per sq ft ah = difference in height above datum plane in, ft ------ = kinetic energy change of one lb of flowing fluid, v equals velocity in ft/sec and g. = 32.174 Ws = Work done by fluid while in flow (similar to shaft work in driving a turbine) W, = Znergy losses due to the irreversibilities of the flowing fluid This equation contains no limiting assumptions and can be made the basis of any fluid flow relationship by limitations or substitutions. The energy units are usually expressed in term* of foot-pounds of energy per pound of mass. Considering the vertical flow of gas through tubing or annu-lus, the external work W, done by the gas is zero. Let us also assume that the value of the kinetic energy function is small and can be neglected. Equation (1) is reduced to: /P: VdP + Ah+Wt = 0 .......(2) Pi The energy loszes PI can be expressed in terms of the well known Fanning equatinn: 4/Lf: rf = —— ............(3) 2g,D where L = the integrated average velocity in ft/sec of vertical flow in a uniform flow string -~h = L v = the integrated average velocity in ft/sec / = dimensionless correlating function

Jan 1, 1951

By Nicholas J. Palethorpe

BACKGROUND Before addressing the above topic in any detail, it is pertinent to provide some background on Australia for those people who have not been there or who have a limited knowledge of our continent. Australia is approximately the size of the continental U.S.A. (excluding Alaska) covering an area of approximately 7 million square kilometres. It has a very dry climate which has led to a concentration of people in major cities scattered around the continents perimeter. Australia's population is currently approximately 15 million. Of the work force of 6.3 million, 72% work in tertiary industries, 20.2% in manufacture, 6.5% in agriculture and 1.3% in mining. Agriculture accounts for 46% of exports, mining for 31% and manufacturing for 23,. Economic activity is based on:- Agriculture This is capital intensive and covers extensive areas and is reliant on low cost farming. A.range of foods and fibres are produced with over 70% of production being exported. Such products are of course subject to the usual vagaries of world market demand and market access, particularly in regard to the operation of trade blocks such as the E.E.C. Mineral Resources A wide range of minerals and an abundance of iron ore, coal, gas and uranium resources provide an increasing volume of exports. However, Autralia's oil resources at present only supply approximately 70% of local requirements. Industry This is diverse and encompasses the processing of agricultural products and mineral ores as well as helping to supply the consumer and capital goods needs of the whole community. However, the small scale of local markets, the geographic isolation of Australia and relatively high wage costs have limited the scale of Australia's industry. Advanced technology has overcome some of these difficulties and high tariff and import quotas protect local producers from the full force of international competition. In 1980/81 Australia's Gross Domestic Product amounted to $130 billion with approximately 24% of this being devoted to investment. In the past decade Australia has experienced inflation rates ranging from 5% to 17% and only modest real growth rates ranging from 0.4% to 5% per annum. We have also experienced, as have many other industrialised nations, continuing wage and price increases. Unemployment presently stands at 6%. MINERAL DEVELOPMENT As the above figures indicate, Australia's continuing prosperity is very dependent on the mineral and agricultrual sectors of our economy. The dependence on these sectors is highlighted by the table below which indicates our position in a global context. [Australian Production of Minerals in a World Context Western World % Total Western World Position Mineral Production 1980 1st Rutile 75 Ilmenite 31 Zircon 70 Bauxite 36 Iron Ore 11 Tungsten 14 2nd Lead (Conc.) 15 3rd Manganese Ore 14 Brown Coal 11 (1979) Zinc (Conc.) 10 4thSilver (Conc.) 10 Nickel 6 5th Tin (Conc.) 5 6th Black Coal 5 Gold 2 Source: Mining Annual Review 1981] The total value of all mineral and fuel exports in 1980/81 is estimated at $5.9 billion. This prominence in the world mineral production has largely been achieved in the last 15 - 20 years. In fact, in the early 1960's, Australia had only 2 major mining export centres, Broken Hill in New South Wales and Mt. Isa in Queensland. This growth is highlighted by the fact that in 1963, for example, we exported only 9 tonnes of iron ore and less than 4 million tonnes of coal. In 1980/81 we exported 75m. tonnes ',of iron ore (valued at $1,117m.) and 47m. tonnes of coal valued at $1,965m.). The growth in our mineral industry has come about because of a number of factors. Abundance of Mineral Resources We are particularly well endowed with mineral resources, as indicated below.

Jan 1, 1982

By Forrest F. Craig

Both early and more recent1 laboratory measurements of gas-oil relative permeabilities were made by subjecting oil-saturated cores to external gas drives. In these runs it was generally assumed that the flowing gas/oil ratio was constant at every point throughout the core. Using this assumption, the resulting equation for the gas-oil relative permeability ratio becomes A where Q is the flow rate at the producing end of the system. This relative permeability was then specified for the average saturation of the core. The assumption and resulting equation fail to recognize that in an external drive, the oil flow rate varies from zero at the injection end to a maximum at the producing end. The assumption also fails to consider the existence of a saturation gradient during an external drive. This saturation gradient is due solely to non-steady-state flow. Some authors2,3,4 have used this erroneous method of calculating relative permeability ratios from external drives. Other authors5,6,7 have used relative permeability ratios in calculating expected field performance by external drive, using Equation (1) and making the incorrect assumption that the saturation is uniform within the driven portion of the reservoir. To find the correct relation between the relative permeability. ratio at the average saturation and the produced gas/oil ratio, the insertion in Equation (1) of a factor, F, to adjust it for oil velocity gradient has been considered. The equation then takes the form Assuming that the oil velocity is directly proportional to the distance from the injection point leads to the conclusion that, for a linear system, the average oil velocity is one-half that at the producing end. Thus, the average gas/oil ratio is twice the produced gas/oil ratio, and Equation (2) is transformed into However, here again the existence of a saturation gradient is not recognized. The produced gas/oil ratio is related to the relative permeability ratio at the saturation of the outlet end of the system. It can be shown quite readily that, if the true log k,/k, saturation relationship is linear over small ranges of saturation, then d log kjko Where aaxv and u2 are the average and outlet saturations, respectively. Examination of Equation (3) shows that the factor F cannot have a single value over the whole saturation range. Another method of dealing with external drives is that of Buckley and Leverett.' Their method, which assumes no gas expansion and outlet end effects, considers fully the saturation gradient present due solely to the displacement mechanism during an external drive, and applies Equation (1) to each point along this gradient. The calculation of production history using relative permeability ratio is straightforward. However, the calculation of relative permeability ratios from production history by this method is impractical. In a recent paper, Welge,1 using an approach similar to that of Buckley and Leverett, presents a simplified method for determining the saturation at the producing end of the system. At thi; saturation, the relative permeability ratio calculated by Equation (1) applies exactly, after breakthrough of the stabilizej front. By this method, relative permeability ratios or production history may be calculated with equal ease. Use of this method as a laboratory tool in determination of relative permeability ratios by external drive requires the knowledge of the total injected fluid volume; in addition to the average saturation and producing ratio. As a preliminary test of this method of calculation, laborntory external gas drives were run on a 3% in. diameter, 11 in. long sample of Nellie Bly consolidated sandstone. This core sample had a permeability of 824 md and a porosity of 28.1 per cent. The driving fluid was air. Three different oils were used, having viscosities of 1.4, 9.8, and 125 cp, respectively. The runs were made under identical pressure differentials,

Jan 1, 1952

By K. T. Aust

The relative interfacial energies of symmetrical tilt boundaries in silver of greater than 99.999 pct purity were measured as a function of orientation difference 0 between 9° and 36° about <001>. The results are in good agreement with the dislocation model of the grain boundary. The energy vs H curve may be represented by the Shockley and Read equation in the form Eoe (0.55—In 8). READ&apos; has indicated that an ideal experimental method of testing the predictions of the dislocation theory of the grain boundary would be to measure the interfacial energies of symmetrical tilt boundaries between face-centered-cubic grains that have a common <001> axis, as a function of orientation difference 0. The misfit H should be defined by a rotation of each grain through an angle of H/2 away from the boundary. Such boundaries might be expected, from geometrical considerations, to be made up of a series of edge dislocations, at least for small values of 0. Energy vs H relationships for tin&apos; and lead:&apos; were previously obtained for an asymmetrical tilt boundary separating two grains in which generally only one grain was rotated about a <001> specimen axis through an angle of H away from the boundary, the other grain remaining fixed. It is possible, however, that the boundary atoms in tin and lead may have attained a symmetrical position, in relation to the orientations of the grains on either side of the boundary, during equilibrium annealing. The two sets of measurements on silicon ferrite&apos; were somewhat better suited for comparison with theory, since the boundaries were symmetrical tilt boundaries in which the axis of relative rotation was <110> in the first series and <l00> in the second; but this alloy has a body-centered-cubic structure. Although the grain boundaries in silver" were symmetrically tilted, the axis of relative rotation was not a simple crystallographic direction, and it varied from boundary to boundary. Although the above measurements were in good agreement with the Shockley and Read equation,"&apos;: it appeared desirable to carry out experiments under the ideal conditions specified by Read.&apos; Consequently, the relative energies of symmetrical tilt boundaries in silver between grains that have a common <001> axis were measured. The technique of equilibrating grain boundaries in oriented tricrystals was used rather than the less satisfactory thermal groove method previously employed for silver bicrystals." Experimental Details and Observations The solidification and seeding techniques, described previouslyh for growing tricrystals of controlled orientations, were used to prepare tricrystal specimens from silver of greater than 99.999 pct purity. The silver was obtained from American Platinum Works, Newark, N. J., in the form of poly crystalline rods 10 cm long and 7 mm in diam. These rods were then cold worked by rolling or drawing them into suitably shaped blanks. Single crystal seeds and tricrystals were grown in a high purity graphite boat by means of a movable furnace with globar heating elements. The graphite boat was enclosed in a Vycor tube, which was first evacuated to about l00/u Hg and then filled with a static argon atmosphere. The maximum temperature outside the Vycor tube was 1250°C. A typical tricrystal specimen of silver is shown in Fig. 1. Grains B and C were grown from seed crystals with a <001> specimen axis or longitudinal direction parallel to the direction of growth and with another <001> direction perpendicular to the surface. Each of the two seed crystals was rotated 18" about the <001> specimen axis but in opposite directions away from the boundary. A symmetrical tilt boundary, therefore, separated grains B and C with an orientation difference 8 of 36" about <001>. Grain A in Fig. 1 was grown from the middle seed with a <110> direction parallel to the specimen axis and with a <001> direction inclined at 45&apos; to the surface of the specimen. The interfaces sepa-

Jan 1, 1957

By U. Andres

U. Andres (The Technion, Haifa, Israel)-The authors are to be congratulated on their research, new and interesting, in a field which, however, the present writer and his assistants have been trying since 1962 to bring to the attention of the scientific and technical world, by means of publications covering several different aspects of this many-sided problem. We announced, in these publications, the results of magneto-hydrostatic (MHD) separation of magnetic and nonmagnetic minerals which we made in magnetic fluids (in paramagnetic solutions and suspensions of magnetic particles) which interact with permanent" and impermanent" inhomogeneous magnetic fields. There is no doubt that the authors have succeeded in creating ferrimagnetic colloids made by dispergation of 10-30 wt % of magnetite in kerosene in the presence of 7-10 vl % oleic acid. It may be judged by [Figs. 3 and 4] that the suspensions represent great stability systems and also "hysteresisless" fluids [(Fig. 5)] which bear witness to the existence of superparamagnetic suspensions. If the viscosity of these colloids is not so high, and the cost of preparing them can be covered, the fluids which have here been received can possibly be of interest to the technology of mineral separation. The potential possibility and technological perspective of similar media is demonstrated in our joint work with G. Bunin18 in 1965. The research carried out on the pushing out of cylinders and spheres from magnetic fluid is impressive; it would, however, have been interesting if the authors had compared their results with those on the pushing out of bodies from magnetic fluids adduced from our work." In my books" there happens to be described an experiment on the pushing out of a sphere in magnetic wedge field, in which the pushing out force of suspension with 5 vl % of magnetite reached 125 g per cc. This testifies to interesting possibilities of these media. Our works&apos; shows also, on the basis of Laplace&apos;s equation the receiving of an interesting distribution of pushing out forces by selection of corresponding configurations of magnetic poles. Objection must be made to the authors&apos; introduction of the term "antigravity force." The direction of this force does not depend on the direction of the gravity force and their opposition, i.e., antigravity is only one particular case. I am in full agreement with the authors&apos; statement that magnetic pushing out forces "can be utilized to separate nonmagnetic objects according to variations in their densities." It is, however, necessary to make only one important addition, which can be found in our research on MHS at the Technion in Haifa, namely, that magnetic fluids in magnetic fields can be utilized in the separation of minerals according to variations in their magnetic susceptibility. The authors&apos; conclusion is especially important considering the fact that in our publications" (including one in English") some designs of MHS separators and some technological results received on MHS-separators have already been described. S. E. Khalafalla and G. W. Reimers (Authors&apos; Reply)We were aware (mostly through Chemical Abstracts of the American Chemical Society) of the use of solutions of paramagnetic salts such as manganous sulfate in agnetohydrodynamic separations by the Russian school headed by G. Bunin."n Professor Andres&apos; contributions to this school are also evident in his application of the ponderomotive forces in an electrolyte placed in electric and magnetic cross-fields to separate coal and other useful minerals." It must be realized, however, that the foregoing research by Bunin et al." involved the use of magnetic fields and field gradients of very large magnitudes, i.e., such that H • grad H - 101 Oe2 per cm. The magnetization of paramagnetic salt solutions range from 0.1 to 0.5 gauss (at fields of a few thousand Oersteds) and never reach saturation. The magnetization of our superparamagnetic ferrofluids ranges from 100 to 500 gauss and reach saturation at fields of a few hundred Oersteds. Hence, to effect the same separation, the quantity "H

Jan 1, 1974

By J. Husson, R. Iffly, M. Gondouin

A systematic variation of well deliverability, as reflected from isochronal back-pressure tests performed at regular intervals, has been observed in some gas condensate wells producing at high rates. The same effects have been obtained using a numerical model of gas and condensate flow which takes into account secondary gasoline deposited in the pore space as a result of pressure reduction, and non - Darcy flow of gas in the vicinity of the wells. Matching calculated values with previous test results has been possible, and future predictions have been obtained. An application of this method to the Hassi Er R&apos;Mel gas-condensate field in Algeria is tentatively shown. INTRODUCTION Flow capacity of gas wells is generally derived from an analysis of back-pressure tests. The empirical equation q = C (?p2)n used by Rawlins and Schellhardtl can be derived rigorously assuming that steady-state radial flow of a dry gas of constant viscosity and compressibility is established during each flow period of the well tests. Furthermore, when Darcy&apos;s law applies in the entire flow region, the theory predicts that the exponent n is equal to 1. In low-permeability reservoirs, it was soon discovered that the time required to reach a stabilized flow often exceeded the duration of the flow periods normally available for testing wells. Consequently, transient gas flow had to be considered instead of the steady-state assumption previously used. This led to the isochronal testing procedure established by Cullender2 which has largely replaced conventional back-pressure testing. For dry gas fields, this method yields definite values of C and n equivalent to those of the empirical equation. These values should remain constant for each well as long as the permeability of the formation and the characteristics of the gas (viscosity and compressibility) do not change appreciably. This is the case when reservoir pressure remains -close to the original value and when the formation near the wellbore remains free of plugging. Under those conditions, stabilized flow potential curves of gas wells can be established from a single sequence of isochronal flow and shut-in periods. An analysis of a pressure build-up following a longer production period provides additional data on the transmissivity (k h/ u) of the reservoir, and eventually on the &apos;drainage radius rd of the well, which can be related to the value of C so that future performance of the well can be predicted using the concepts developed by A. Houpeurt.3 At high flow rates, Darcy&apos;s law no longer applies in the vicinity of the wellbore, and inertial effects in the high velocity gas flow introduce additional pressure drops. As a consequence, exponent n of the backpressure tests becomes smaller than 1, and a slight curvature of the log-log plot of ?p2 vs q can be predicted when going from very low to high rates of flow (Elenbaas and Katz).5 The effects of variations of viscosity and compressibility with pressure on the radial flow of dry gas in an infinite reservoir were taken into account by Jenkins and Aronofsky.7 Numerical solutions of the transient flow of an ideal gas in finite radial reservoirs were presented by Bruce, Peaceman, Rachford and Rice.8 In the case of gas condensate wells, however, the presence of gasoline in the pore space as soon as reservoir pressure is reduced below the dewpoint pressure further complicates the interpretation of flow tests so that the prediction of stabilized well performance becomes very difficult. Field observation shows that both C and n derived from isochronal tests vary in time, even when reservoir pressure has not changed appreciably. Such a variation cannot be attributed to any change of the gas characteristics, and must result from the effect of a gasoline saturation on the gas flow.

By R. Schuhmann

D. W. Fuerstenau (Associate Professor of Metallurgy, Dept. of Mineral Technology, University of California, Berkeley, Calif.) In his excellent paper, Dr. Schuhmann has proposed a mechanism for conlrninution which can be utilized to explain a number of phenomena which have been observed in the size reduction of solids. The difference between the product obtained from a ball mill and from a rod mill can readily be explained by Dr. Schuhmann&apos;s hypothesis. Consider that the size distribution of the product from each comminution event is given by y; - 100 (x/ki) (1) where Yi is the cumulative percent finer than size x, ki is the size modulus, and is the distribution modulus for the single comminution event. The total product from a cornminution operation will be the summation of the products from all of the individual cornrninution events. If the size moduli from the various comminution events range from kl to kn , the size distribution of the total product will be a straight line on a log-log plot only to a particle size equal to k ,. Considering 100 volurnes of feed, the value of y for x equal to kl is Riven by where y is the cumulative percent finer than size kj, z is the number of comminution events of kind i, and vi is the volume of each particle crushed. y=z1 qv1+z2 qv2 (k1/k2) +.............zn qvn (k1/kn) (3) F or particle sizes greater than k1 7 e. 6. x = k3, the value of y will be given by k3 A y = z1 qv1 + z2bv2 + Z3uV3 + zn zn qvn (k3/kn)(4) Only when x is equal to k, will y equal 100. &apos;Thus, the total product from a comminution operation will be a straight line on a log-log plot strictly up to a particle size equal to kl and above that size the line will curve so that y will equal 100 for the coarsest particle or k, (see Fig. 1). The actual shape of the size of distribution curve will depend upon the relative number of conmninution events of different kinds. In the case of ball milling, the ratio kn/kl may be pite large and there will be considerable deviation at the coarse sizes in a Schuhmann plot. kIowever, the grinding action in a rod mill is such that fine particles are protected from the rods by the coarse particles. Consequently, kn/kl will be quite small and the product from a rod mill will give a straight line Schuhmann plot almost to y equal 100%. In his paper, Dr. Schuhmann also has presented a very straightforward definition of grindability derived from the basic energy equation E = Ak-a (5) where E is the energy required to comminute a material to a size modulus of k and A is a constant. According to Dr. Schuhmann, the grindability of a material is the volume of material comminuted finer than unit size per unit of energy expended, this being 1/A. For comparison of grindabilities at any size two numbers are required, both = and 1/A, and anyone tabulating grindabilities should give both of these numbers. Concerning the experimental determination of grindabilities, if the material being comminuted possesses distinct cleavage, the energy re quired for comminution is given by E = Ak-(a-y) (6) where y is small number less than = . Because of cleavage, less energy will be required for comminution. To be certain that cleavage does not affect the gindability of the material being tested, it is suggested that the method used by Dr. Charles for determining grindability be used to ascertain if the deviation exponent y is zero or finite. R. Schuhmann, Jr. (Author&apos;s Reply) Dr. Fuerstenau&apos;s interesting comparison of rod milling and ball milling

Jan 1, 1961

By R. Simon

The sources of energy dissipation for concentrated loadings on rock are considered in an attempt to account for the experimentally measured magnitude of the work required to break out a unit volume of rock from the free surface of an essentially semi-infinite medium. It is concluded that most of this work probably represents the elastic strain energy developed by the loading in a much larger volume of rock beneath the loaded region than the volume of the rock fragment broken out to the side of the loaded region. This strain energy is largely dissipated in the form of stress waves generated by the high rate of unloading produced by the propagating cracks. The energies associated with the formation of the new surfaces of the cracks and with the stress waves generated directly by the loading process are computed to be negligibly small. Possibilities for improving the utilization of energy to drill rod. subject to the geometrical limitations imposed by down-hole operation, are discussed. It is pointed out that any such possible improvements would probably have to be differential ones, since each rock configuration of more favorabIe loading geometry that can be created down the hole is accompanied by a complementary configuration of less favorable loading geometry. INTRODUCTION Dislodging each cubic inch of rock from the bottom of the hole by the action of a bit requires the expenditure of an amount of energy that varies from approximately 5,000 in.-lb to approximately 100,000 in.-lb, depending on the hardness of the rock, or, more technically, upon its fragmentation strength.1 In this paper we will discuss (I) why the energy expended in drilling is so large, (2) what happens to this energy upon completion of the drilling process, and (3) what are the possibilities for reducing the magnitude of the energy required to drill rock. DETERMINATION OF ROCK DRILLING ENERGY The volume of rock removed per unit time from the bottom of a hole of diameter D is evidently (7/4)D2R, where R is the rate of penetration of the bit. If P is the rate at which work is done by the bit on the rock at the bottom of the hole, the energy required to break out a unit volume of rock is given by: Ev =(4/p) P/D2R............(i) For rotary drilling, of either the rolling-cone or drag-bit varicty, P = 2pLN, where L is the torque resistance to rotation at the bottom of the hole and N is the rate of rotation of the bit. L is essentially the same as the torque measured at the rotary table only when drilling in shallow holes. The energies expended in rotating the drill string against the frictional resistance of the walls of the hole and and against the viscous drag of the drilling fluid are extraneous to the subject under consideration, although these may be much greater in magnitude than E, when drilling in a deep hole. For percussion drilling, P = fE where f is the percussion frequency and E is the work done on the rock per impact. The latter quantity can be both computed and measured for a percussion drilling machine. 2 (Under satisfactory drilling conditions, defined in terms of ranges of numerical values of certain dimensionless parameters, E is only 30 to 50 per cent less than the impact energy of the striker.2) Alternatively, E, may be measured directly by dropping chisels shaped like bit edges, backed by rigid weights, onto the surfaces of laboratory rock samples of effectively semi-infinite extent. Under these circumstances, essentially all of the impact energy is converted to work done on the rock, and the relationships among volume of rock broken out, chisel shape, impact energy and indexing distances can be obtained.3 The values of the energy required to break out a unit volume of rock under favorable circumstances are substantially in agreement for rotary drilling, percussion drilling and drop testing at atmospheric pressure. The energy per unit volume is a quantity of the order of magnitude of roughly twice the com-pressive strength of the rock as measured by a uniaxial loading test. The phrase "order of magnitude" in this paper means from about 1/3 as much to 3 times as much; i.e., the energy per unit volume may range from roughly the same up to several times

By R. Schuhmann

Distribution of material in the fine sizes of a comminution product generally is well represented by the empirical equation&apos; 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 S. F. Shaw

IT IS generally recognized that the natural gas absorbed in petroleum plays the leading r6le in moving the oil through the sands to the well and supplies the energy that delivers the oil to the surface when the wells are flowing. It is the purpose of this paper to note the amount of oil that could be lifted by this gas if 100 per cent. efficiency were obtained and the efficiencies actually obtained in the use of this gas. An article in MINING AND METALLURGY,&apos; stated that in a Wyoming field, with a pressure of 300 lb. per sq. in., there would be absorbed 65 cu. ft. of gas per barrel, or 1.55 cu. ft. per gal. of oil. The writer does not know if data have been obtained, by experiment, for pressures beyond those indicated in this article, therefore, for the purpose of illustration, he has projected this curve to obtain figures for pressures greater than 300 lb., although probably experiment will show that the extended line is not quite correct.

Jan 1, 1926

By T. H. Weldon, L. V. Whiton, R. R. McNaughton, J. H. Hargrave

Oxygen enriched air is being used quite extensively in the metallurgical plants of The Consolidated Mining and Smelting Co. of Canada, Limited, at Trail, B.C. The oxygen used for this purpose is a by-product from the Company&apos;s chemical plants located in the area. Most of the ore treated in the Trail metallurgical plants comes from Com-inco&apos;s Sullivan Mine at Kimberley, B.C. At Kimberley the ore is milled to produce lead and zinc concentrates which are shipped to Trail for further treatment to metal. One section of this paper deals with the use of oxygen enriched air in the suspension roasting of the zinc concentrates. In this process the concentrate is calcined for leaching preparatory to electrolytic recovery of the zinc. A second section of the paper describes the use of oxygen enriched air in operations at the lead smelter. There oxygen enriched air is used in the blast to the lead blast furnaces and in the slag fuming furnace which recovers the lead and zinc contained in lead blast furnace slag. The final section of the paper outlines the precautions necessary for the safe use of oxygen enriched air in any plant operation. The Use of Oxygen Enriched Air in the Suspension Zinc Roasters The suspension roasting of zinc concentrate developed at Trail, B.C., has been described in AIME Vol. 121, "The Electrolytic Zinc Plant of The Consolidated Mining and Smelting Company of Canada, Limited" by B. A. Stimmel, W. 11. Hannay and K. D. McBean. Since the publication of that paper, the use of oxygen enriched air in suspension roasting has been introduced as regular practice with marked advantage. The relative importance of a specific advantage may vary with changing conditions, but, in general, it may be stated that improved operation has been achieved at increased capacities. Suspension roasting is carried out at Trail in converted standard 25 ft diam Wedge roasters. The 2nd, 3rd, 4th and 5th roasting hearths have been removed and the drying hearth covered over. Drying of the concentrate is done on the drying hearth and the 1st roasting hearth. The dried concentrate, after any lump&apos;s have been broken up in a ball mill, is fed to a single burner located in the upper part of the combustion chamber. Oxygen is introduced at the burner fan along with the gases from drying, returned combustion gases from the waste heat boiler outlet and the required amount of new air. Up to 60 pct of the concentrate settles out on the 6th roasting hearth, the rest passilig out of the roaster. The product collected in the waste heat boiler is finished calcine, but the dust collected in the cyclones after the boilers is returned to the base of the combustion chamber where the sulphate is decomposed. Gases from the c&apos;yclones go to a Cottrell precipitator. The discharges from the 7th roasting hearth and the waste heat boiler are combined with the Cottrell dust to give the finished calcine. Following small scale tests started in 1933, oxygen enriched air has been used continuously in the suspension roasting of zinc concentrate in Trail, beginning in 1937. A significant factor in promoting its use in this operation was the availability of by-product oxygen from the Company&apos;s near-by Chemical and Fertilizer Division. To-day it is standard practice at Trail to use oxygen enriched air for zinc concentrate roasting. The most important requirement in roasting a zinc concentrate for an electrolytic plant is that the zinc in the calcine should have maximum solubility. It is also desirable at Trail that the gas produced for the manufacture of suhhuric acid should have a ~naximum concentration of SO2 and that a substantial recovery of waste heat from the gas be achieved. High solubility of zinc requires that the sulphide sulphur and zinc ferrite in the calcine be kept low. These are both functions of temperature and time, with formation of zinc ferrite also dependent on contact between the iron and zinc particles. One of the inherent advantages of suspension roasting is that minimum time and contact are achieved. The limit on temperature is imposed by the fusion point of the concentrate and not by the need to control zinc ferrite formation. Operating temperatures are normally maintained within the limits of 1725" and 1850°F. A relatively low zinc sulphate in the calcine is required at Trail, and this results from discharging the calcine from the high sulphur dioxide atmosphere at a temperature above 1600°F.

Jan 1, 1950

A little more than a year and a half ago construction started on Canada&apos;s largest initial milling plant. Today the zinc concentrating plant 7 miles north of Barraute sits on ground that once was forest and clay flats. The plant is part of Barvue Mines Ltd. zinc-silver operation. Employees of the plant live in the model community of Barvue, built like the plant in record time. About 5000 tons of ore are handled by the plant daily, with an output of 200 tons of zinc concentrate. Management hopes to raise capacity of the $8 million plant to 6000 tons.

Jan 8, 1953

The 38 companies listed below are those substantially engaged in the mining of metals, non- metals, or coal which appeared among the country&apos;s top 500 (on the basis of sales volume) listed in the July 1966 issue of Fortune. General Motors, with $20.7 billion in sales, heads the entire list, and a prominent member of the mining industry, Island Creek Coal, occupies the 500th spot with sales of $109.6 million. The lists of "industry medians" following mining&apos;s "Top 38" reveal how well the mining industry fared in 1965 in various salient categories of business performance.

Jan 9, 1966

By O. E. KEOUGH

AT the Tooele custom lead-zinc ore concentrator,&apos; two sections, each having a daily capacity of 500 to 600 tons, are operated on slightly different types of ores with but little difference in flow sheet and reagents. The ores comprising the two general classes are the coarsely crystalline type, with low pyrite and sphalerite low in iron content, from the Park City district, and the more finely disseminated higher pyrite ores from Bingham, which are more difficult to treat by flotation. Laboratory testing determines with which general class a new ore will work better. The ores are crushed to minus 1 1/4-in. size at either the Utah Ore Sampling Company&apos;s plant at Murray, Utah, the Tooele smelter sampling plant, or the crushing plant at the concentrator. The ores for each section are mixed to maintain as nearly as possible uniform lead and zinc contents.

Jan 1, 1930

By Thomas Henderson

Strategically, intermaterials competition is important to companies involved in raw materials extraction, smelting, refining or other processing, semi-finished and finished goods fabrication, and end-use consumers. In other words, just about every company involved in manufacturing. The array of materials covered by other papers in this session is testimony to the subject&apos;s breadth. In trying to provide an over- view of such a broad topic, the purpose has been to relate some of the common characteristics of the markets, economics and technology of raw materials, and to describe some of the ways in which Intermaterials competition impacts the markets, economics and technology of raw materials.

Jan 1, 1982