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By Kenneth M. Bravinder, Jonathan E. Koogle, Dean H. Sheldon

THE solution of subsurface geological problems requires an analysis of vertical and horizontal dimensional relationships. For many, the ability to visualize structures in three dimensions is not easily acquired, and for areas where the structural conditions are complicated by faulting the need arises for illustrative models. This has led to the many forms of peg models, in which the usual course of construction includes a peg for the well bore, and the suspension of strings from peg to peg to illustrate the formation correlation between wells, thus developing the usual folds and flexures. Such models proved unsatisfactory for the analysis of fault problems. Consideration was then given to the use of transparent materials. A model of the Dominguez oil field was constructed in 1935, which consisted of a series of glass panels from well to well on which the formations were shown in color. This model, though serving its purpose at the time, proved to be impractical from the standpoints of cost, breakage hazard, and extreme difficulty in maintenance. The cost of materials was estimated to be approximately $300, and the construction required the services of one man for a period of from 10 to 12 months. Maintenance after actual construction was an additional expense. Consideration was then given to the use of celluloid, and after some experimental work it was found that the advantages in its use were: (I) a short construction time, (2) ease of handling the material, (3) sections are easily reworked, (4) low cost. CONSTRUCTION It would be of little value to recount here the difficulties encountered in the building of the first celluloid model. The conclusions drawn, however, suggest the following procedure for those who contemplate the construction of such a model: I. Determination of the Scale.-Accurate geological interpretation requires the use of the same scale horizontally and vertically. It should be emphasized that the limited size of the celluloid sheets (20 by 48 in. maximum dimensions), the area to be studied, the maximum depths to be plotted, and area of possible future development are the controlling factors in adopting a scale to be used. 2. Construction of the Base.-A satisfactory base may be constructed by using two pieces of three-ply wood separated by strips of i by 2-in. material. This type of construction prevents warping of the base. The size of the base is determined by the dimensions of the area to be studied plus 5 in. on all sides. A map of the same scale as that adopted for the model, indicating the lines of section to be used, should be fastened to the base. 3. Construction of the Frame (Fig. I).-In the construction of the frame the use of 1/2-in. o.d. pipe welded in a box shape is suggested, with two horizontal rods on each of the four sides, top and bottom. The top rod should be at least I in. higher than the

Jan 1, 1941

By C. J. Sparks, J. C. Ogle, E. A. Starke

DEFORMATION of cold-rolled fcc metals and alloys produces one or the other of two types of rolling textures, usually referred to as the copper type or the brass type. The pure fcc metals, with the exception of silver, develop the copper-type texture when rolled at room temperature. Silver, brass, and many other alloys develop the brass-type texture. Transition from one texture type to another can also occur. Solute additions to pure copper cause a transition from the copper-type to the brass-type texture.' Varying the deformation temperature also produces texture transitions. Copper rolled at -196°C developed the brass-type texture2 and silver rolled at elevated temperatures, 200°C, developed the copper-type texture.374 The work of Haessner,' Smallman and Green,6 and Hu and coworkers3'4 seems to have established an apparent correlation between the type of rolling texture (and the texture transition) and the stacking-fault probability. Thus materials with low stacking-fault energies are thought to develop the brass-type texture while high stacking-fault energy materials develop the copper-type texture. Mikkola and cohen7 measured by X-ray diffraction techniques the stacking-fault probability, a, in the ordered and disordered fcc alloy Cu3Au. Unexpectedly, ordered specimens showed about twice the stacking-fault probability as the disordered ones. Marcinkowski and Zwe118 also found stacking faults in ordered Cu,Au, but not in the disordered CusAu. More recently, Camanzi and coworkers9 have shown that the intrinsic stacking-fault energy of Cu3Au undergoes a considerable decrease (approximately 100 ergs per sq cm) when passing from the disordered to the ordered phase. In general, deformation reduces order'' but some of our measurements showed that some long-range order still remains in Cu3Au after cold reduction by rolling 80 pct. From the above considerations Cu3Au offers the possibility of studying the texture transition in fcc alloys without varying either the temperature or the composition. If the stacking-fault energy correlation is correct, an initially ordered alloy should develop the brass-type texture during rolling while a disordered alloy should develop the copper-type texture. The present communication describes some initial results of experiments designed to investigate this possibility. The rolling textures of ordered and disordered Cu3Au samples were determined after various amounts of deformation. The ordered sample had an initial long-range-order parameter of S = 1, determined from X-ray diffraction. The grain size of both the ordered and disordered samples was the same, and the initial textures of both were also identical and approximately random. No obvious differences in the deformation textures of the ordered and disordered materials were detected up to 35 pct reduction and the (111) pole figures of Fig. 1 resemble that of pure copper. However, with reduction larger than 35 pct and less than 50 pct, a marked deviation from the copper texture was noted in the ordered material. The newly developed orientation was close to the (110)[i12] ideal orientations re-

Jan 1, 1970

By J. B. Newkirk, W. G. Martin

Oxidized cobalt powder is known to have a magnetic hysteresis loop which is asymmetric with respect to the magnetization axis. The experiment described herein shows that the orientation relationship between the basal plane of hexagonal cobalt and the oxide which forms upon it at 400°C is {111}ux//{100.1 }co <110>ux//<11.O> This orientation relationship allows a magnetic interaction between the antiferromagnetic oxide and the ferromagnetic substrate which could account for the offset hysteresis loop. Oxidized nickel powder has a symmetrical hysteresis loop and so is apparently not influenced by any magnetic interaction between metal and oxide. The orientation of the oxide (cubic) was found to be identical with that of the nickel substrate when the oxide forms on a polished surface parallel with {111} Ni. IN 1956 W. H. Meiklejohn and C. P. Bean discovered that fine particles of cobalt which had been prepared in a certain way have a magnetic hysteresis loop which in a strong field is asymmetric relative to the magnetization axis.&apos; The cobalt powder exhibited this unusual magnetic property only after it had been oxidized in air or oxygen; it lost the shifted hysteresis loop when the oxide was reduced in hydrogen. X-ray and neutron diffraction&apos; experiments showed the presence of cobaltous oxide (COO) and hexagonal cobalt in samples exhibiting the biased hysteresis loop and specifically showed no indication of any other compound with the exception of mercuric oxide.* The magnetic hysteresis loop becomes symmetrical above the Nee1 temperature (paramagnetic state) of COO. Therefore, it was concluded that the anomalous magnetic behavior is associated with the influence of cobaltous oxide upon the metallic cobalt. It has been proposed that crystallographic coherency may exist between the cobalt and a plane of some antiferromagnetic material which has unbalanced spin distribution and sufficient magnetic anisotropy to hold its spin in the direction which existed when the specimen was cooled. COO has these properties. Therefore, Roth&apos; has suggested that cobaltous oxide may form with a {111} plane parallel and coherent with the basal plane of the hexagonal cobalt metal and proposed that, as consequence of the antiferromagnetic interaction between COO and the underlying cobalt, the magnetization direction in oxidized fine Co particles may be ro-lated from the easy c-direction.2 Such a relationship, he proposed, would explain the observed magnetic effect in oxidized cobalt powder. The main purpose of this study was to determine the orientation relationship, if any, between the basal plane of cobalt and the oxide which forms upon it. Attempts to produce a film of COO which was strong enough to be handled were not successful. However we did succeed in making a film of CoCo2O, which was strong enough to be removed from the cobalt substrate and mounted on an electron-microscope grid. Because of the close structural similarity of COO and CoCo2O, we believe the orientation relation found for CoCo2O, on cobalt probably also holds for COO on cobalt. The epitaxial relationship of NiO and Ni also was investigated. To date no shifted hysteresis loop has been observed with nickel powder. However, the similarity of atomic array in cobalt and nickel leads to the prediction that a {111} of NiO may be parallel with a (111} of nickel. Experimental Method Cobalt-Cobalt Oxide—A coarse-grained specimen of a (hexagonal) cobalt was prepared by allowing a large crystal of fee cobalt to transform slowly at 400°C. The crystal was then cut to expose a surface parallel with the basal plane. A back reflection Laue X-ray photograph showed that the 00.1 plane was within 2" of the cut surface. The surface was mechanically polished and then electro polished in 85 pet orthophosphoric acid after which the crystal was held for 30 min at 400°C in air. During this heat-treatment the surface darkened slightly due to the oxide film which formed. The film was not thick enough to give an X-ray diffraction pattern, even by a glancing-angle technique. Glancing-angle electron diffraction was not possible either, owing to the interference of the electron beam with the high magnetic fields which exist at the 00.1 surface of the cobalt. However, it was possible to make an electron-diffraction photograph of the oxide film by stripping it from the cobalt substrate using the method described later. By maintaining reference marks carefully, it was possible to preserve the orientation relationship between the stripped film and the substrate on which it was formed. The oxidized surface of the crystal was first covered with a 1 pet solution of collodion (cellulose nitrate) in amyl acetate. When the film was dry, small rectangles were scored on the surface with a needle point. The specimen was then repolished

Jan 1, 1959

By Russell Franks, W. O. Binder

In recent years a considerable tonnage of the 18 per cent chromium and 8 per cent nickel steels has been used in lightweight high-strength structures that must be resistant to deterioration under atmospheric conditions. These steels are austenitic in character and their strength can best be increased by cold-working. Although a marked increase in tensile strength is obtained, claims have been made that the cold-worked steels do not possess fully satisfactory elastic propertics. Under certain conditions, however, this has caused no great concern, because, as explained by Ragsdale,1,2 he cold-rolled steels have been widely applied for construction of lightweight, high-strength trains using spot welding for joining purposes. Ffield3 agrees that the ease with which thin sections of the cold-rolled 18-8 steels can be joined by spot welding is greatly in their favor, but that they would be more satisfactory if higher elastic properties and a higher modulus of elasticity could be obtained. He states that, as far as the mechanical properties are concerned, the straight chromium stainless steel is superior to the cold-worked 18-8 steel, and in order to illustrate, he prcsented the data of Fig. 1, which also give a stress-strain curve for a low-carbon 13 per cent Cr steel. Ffield† proposed that columbium-bearing or titanium-bearing cold-rolled 18-8 steels be given a heat-treatment (preferably for 1 hr.) in the range 400" to 600°C. to precipitate carbides in order to obtain better mechanical properties. Discussions by Sutton,4,5 de Ganahl,6,7,8 Krivobok and Lincoln,&apos; Knerr,l0 and Thaden11 have given further information on the properties of the cold-rolled 18-8 steels compared with other materials used in lightweight, high-strength structures. These investigators and others point out that the ultimate tensile strength of the cold-rolled 18-8 steel can be increased to 200,000 lb. per sq. in. or higher; Krivobok et al. indicate that

Jan 1, 1940

By L. D. Clark

Pay-out, the ratio of capital expenditure to annual cash return can be very revealing when employed to gage the merits of a proposed mining venture. It is a quantity, numerically but not physically equal to a factor, by which cash return may be discounted to equal required investment. The fewer the pay-out years, the greater the average annual earning power of the project. Considerable material has been written in the past about the most important and difficult problem mining management faces in evaluation of a property: the optimum production rate for the venture. What daily capacity should be adopted for the proposed operation that it may be the most rewarding to those underwriting it? H.C. Hoover has said, "The most important objective is the least cost per ton mined; and minimum working costs can only be gained by the most intensive production." Theoretically and practically, the quicker the ore is removed, the more profitable should be the venture because of the saving in fixed charges, the consequent lower unit costs and the interest unearned by the profit from the unmined ore. Yet, such performance will be tempered by policy and a restraint imposed by the ratio of capital expenditure to annual cash return. * The &apos;Annual net profit after taxes but including depreciation and depletion allowances. former will be based upon an analysis of the economic climate anticipated, with its many ramifications including the demand for the product, prices, tax laws, the conduct of labor and related wage scales. The variables and intangibles involved in estimating the optimum production rate for a mine do not readily lend themselves to mathematical relationships or equations and the problem can become complex indeed. The more readily applicable and relatively faster appraisal method will be used which, in addition to being reasonably adequate in most cases, will serve as a guide, should a more expanded and detailed estimate be desired. The first question is what is company policy with respect to the rate of investment return? This will depend on that significant period of time referred to as the pay-out, which can be defined as the estimated interval during which capital investment is returned through application of annual cash return to its recovery. Expressed another way, it is the ratio of capital expenditure to the annual cash return. For example, if the capital investment is $C and the annual cash return is $A (after Federal tax), then $C/$A = the pay-out period in years. That is to say, if $C = $1,500,000 and $A = $214,000, then which is the mathematical expression for the period of time it takes to recover the funds originally invested in a mining enterprise. The relationship yp = $C/$A can be useful — an expression which requires only the significant tern yp to be fixed; in other words, the long range investment policy. For what is the pay-out period but an expression of that policy in terms of the speed, within practical limits, with which an individual or company may want to recover capital expenditure? A policy by which pay-out will be fixed is imperative. This ratio must lie within relatively narrow limits governed by what is economically feasible. It cannot be too low, rendering the ratio impractical or absurd, nor too great, delaying the return of capital for reinvestment beyond a commensurate interval. What pay-out will govern this policy? Federal corporate taxes would seem to present the best measure. Consider mining in Canada where new mines are exempt from Federal income tax for 3½ yr from date of first production. This exemption expressly allows The Act states: "A corporation . . . i s not required to include the profits . . . for the period of 36 months commencing with the day on which the mine came into production in reasonable commercial quantites." a company to recover capital expenditure during the early stages of production. With such an incentive, it is logical to strive for as much capital redemption as possible during this period. Thus, 3½ yr would be a major contribution to the pay-out in Canada. Furthermore, it generally takes from 1½ to 3 yr to develop

Jan 1, 1963

By K. I. Savage, L. G. Austin, S. C. Sun

The comminution environment - the physical, chemical, and solid state - was studied for silicon carbide ground wet in a laboratory vibratory mill. The relative importance of these variables on the grinding rate of producing a material 50% passing 5µ from a powder feed of 50% passing 19µ was established.

Jan 1, 1975

By R. G. Clay, C. F. Seaman

THE mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that time, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered, the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. ORIGINAL EQUIPMENT AND INSTALLATION A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots and the up position would connect it to the line from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1942

By Wayne Loel

The application of aerial photographs to all phases of geologic mapping is set forth, indicating the advantages to be gained in different types of country and undcr varying climatic conditions. Methods of field and office procedure are given from a geological point of view, as well as suggestions regarding the best methods of aerial mapping for different uses and under varying field conditions, including the study of submarinc geology. Aerial photographs with accompanying geologic maps are shown to illustrate the appearance from the air of different geologic features under varying climatic and topographic conditions. Introduction The following paper is in no sense a contribution to knowledge; it is merely an attempt to indicate some of the benefits that may be expected from the use of aerial photographs in geologic mapping and to mention some inherent limitations. The writer has used aerial photographs in his own work for the past 13 years and has had ample opportunity to observe their use in the hands of others. From this experience certain conclusions have been drawn, which may be helpful to those using aerial photographs for the first time. It must he remembered that a photograph of the earth&apos;s surface can delineate only that which the eye can see; therefore the geologist should not expect more from an aerial photograph than he would be able to observe were he in the photographer&apos;s airplane looking at the area to be photographed. Theoretically, if the geologist had sufficient artistic ability, he could draw as good a geologic map from his position in the plane as he could from observation of the photograph. Practically this cannot be done because the airplane moves at a high rate of speed, which in itself prohibits detailed observation and, no matter how great the geologist&apos;s artistic ability, he could not reproduce the detail that can be recorded by a camera. An aerial photograph is a picture map of an area. The picture may be studied in the office until the geologist is thoroughly familiar with every minute detail of the surface of the ground he is about to geologize. If it is an area, underlain by sedimentary rocks, much may be learned

Jan 1, 1941

By Frederick J. Pope

The protaxis of the North American continent consists, as is well known, of a large V-shaped area of Archæan rocks, which lie for the most part in the Dominion of Canada, and occupy an area of not less than 2,031,000 square miles. The western limit of this area extends from &apos;Lake Superior in a northwesterly direction to the Arctic Ocean, the southern is found along the Lower St. Lawrence as far as Labrador. A minor area, spoken of as the Adirondack region of northern New York, has a somewhat broken connection with the main body in a narrow band of igneous rock, which crosses the Upper St. Lawrence where it leaves Lake Ontario. The ancient crystallines are also presumably beneath all the later sediments, hut the above brief description outlines their principal exposures. This great rock-complex consists, for the most part, of a series of gneisses, which present great variations in both structure and composition. In certain sections areas of gneisses of the general composition of granite, and remarkably uniform in character, prevail. They consist largely, if not exclusively, of crystalline rocks in which a banding or foliation has been induced by pressure. These rocks constitute the basal member of the Laurentian system, and by later writers have been referred to as the Ottawa gneiss. In other and quite extended areas, gneisses, regarded by many as composed of altered sediments, are associated with crystalline limestones, quartzites, and much injected igneous matter. To this series the term Grenville has been applied. Both of these series of gneisses are penetrated by various igneous masses, of which the most important are great intrusioiis of several members of the gabbro family, such as anortho-sites, norites, pyroxenites, and gabbros proper; the whole form-

Jan 1, 1900

By John J. Reed

This paper was prepared as a supplement to the general review by Dr. N. G. W. Cook. His analysis of the problem in terms of the strain energy involved in making the excavation is excellent. It will contribute much towards a clearer understanding of rock behavior and control. Several of the practical applications described herein can be directly related to Dr. Cook&apos;s theories and conclusions, and are based on the very same principles. CONTROLLED BLASTING IN ROCK EXCAVATION Generally speaking, the natural rock is a relatively strong and competent structural material, if we conserve its inherent good properties and perhaps reinforce its weak spots. The difficulty often lies in blasting out the tunnel or chamber without damaging the walls and roof we wish to have remain afterwards. Much progress has been made in the last five years in better understanding the controlled excavation of rock. The specialized techniques known as smooth wall blasting and presplit blasting have now developed from the status of curiosities of rock excavation to the point where these techniques are specified in contracts drawn up for such rock excavation projects as tunnels, underground pourer houses, etc. These techniques have been well described and discussed in previous publications; there is no need to reiterate here. It might be well to point out that recent research has demonstrated the effect that inherent rock stress may have on presplitting in certain directions. Material can be presplit most readily parallel to the direction of maximum compressive stress, and this may not always be in the direction desired. Inherent rock stress measurements by stress relief over coring, etc., may have important value in deciding how to best presplit proposed excavations. In addition, recent work at the Colorado School of Mines has demonstrated the feasibility of con-

Jan 1, 1967

By Ronald J. Roman, Craig Olsen

Alterations were performed on a previously developed heap leaching model representing sulfuric acid solution leaching of a malachite sandstone ore. The alterations were made in order to approach a more realistic solution to the leaching equations used in the model and to provide a means of obtaining a constant acid con- centration in the model&apos; s simulated heap effluent solutions. This constant acid concentration is desired if the effluent solutions are to be treated in a solvent extraction-electrowinning plant. An economic study was made showing a use of the leaching model for an imaginary heap leach operation. The example illustrates a method of selecting heap leaching parameters if data are available on mining costs, dump costs, capital cost, etc. This was achieved using an incremental investment technique.

Jan 1, 1974

By Margaret Dienes, Karl L. Fetters

It is recognized that ingot structure is important in determining the quality of finished steel. Such elements of ingot structure as the size, shape and distribution of primary crystals; the size, shape and distribution of shrinkage cavities; the degree and kind of segregation (both macro and micro); the distribution of nonmetallic impurities and gas cavities are each the result of solidification phenomena. &apos;These structural features vary with variations in ingot practice. Their control requires a knowledge of their correlation with the variables of ingot practice. The structures of full-sized commercial steel ingots have been studied extensively by the Steel Ingots Committee of the Iron and Steel Institute.l This work has contributed much to our knowledge of ingot structure and the factors that control it. These investigations of necessity have been limited in scope because of their costliness and the great amount of time required. Studies of the structure of small steel ingots have not been very satisfactory because of the rapid rate of freezing of the ingots. Their structures are thus difficult to resolve and to interpret. Medium sizes of steel ingots in the ranges of 150 to 500 lb. do offer opportunity for study, but their cost also becomes greater and their use less expeditious. Experimental models for the study of steel-ingot structure have been made of nonferrous metals. Gathmann2 used lead-tin alloys and reported that they were satisfactory for determination of the location of the major shrinkage cavities and the Liplanes of weaknessv resulting from the pattern of crystallization. Certain features of mold design have been illustrated by these results. Bezdenezhnihk3 has used copper and aluminum ingots similarly and has shown the effect of mold-wall thickness on the depth of columnar grains. These studies of nonferrous models have been useful, but their results have been limited because the analogy with steel ingots obviously is not a close one. Model wax ingots have also been useful in pointing out important features of mold design. Gathmann4 watched the solidification of wax ingots, through a glass-walled mold, and has reported the effects of mold type on the location of the shrinkage cavities. Segregation was simulated by using a mixture of paraffin and stearine and adding a red dye that on freezing segregated preferentially to the stearine. Brearley5 also used wax models to determine the effects of some of the principal features of mold design. Carlsson and Hultgren6 dsed stearic acid to show that the columnar crystals of ingots tend to slope upward progressively from the mold wall toward the center of the ingot. As with the nonferrous ingots, the wax ingots have

Jan 1, 1943

By O. J. Malacarne, D. G. Millenbruch

The use of a series of thickeners for washing of the slime fraction of a leached ore is not uncommon in the uranium industry; however, few ore processing mills have used a thickener circuit for washing leached total ore. The common practice is to wash the sand fraction in classifiers and the slime fraction only in thickeners. At the Uravan mill of Union Carbide Nuclear Co., a total ore acid leach is employed to extract uranium values from sandstone ores. These leached ores are then washed at a nominal rate of from 1000 to 1300 tpd in a series of washing thickeners. A typical screen analysis of the ore being washed in these thickeners is: 25 pct, -200 mesh; 3 pct, +35 mesh; and about 1 pct, +14 mesh. The amount of slimes (-200 mesh) will vary from 20 to 30 pct as the mill receives ore from 150 different mines. The washing liquor is acidic, from a pH of about 2 in the first thickener to 0.6 pH in the last thickener and, of course, all acid duty equipment must be used. The Uravan mill of Union Carbide Nuclear Co. is located in Montrose County, Colo. and is physically located on a canyon wall along the San Miguel River. The thickeners referred to in this paper are located on top of the mesa above the canyon. For purposes of identification, these thickeners are numbered 1 through 8 from south to north. GENERAL DESCRIPTION OF CIRCUIT The circuit utilized consists of eight wooden fir tanks, 60-ft diam by 12-ft deep, set in a line and staggered in elevation by approximately 3 ft. The washing water flows through the thickeners by gravity, while the ore pulp is pumped uphill from thickener to thickener. Adequate piping and pumping capacity are available so that any of the thickeners can be bypassed in the circuit. All the controls for the

Jan 1, 1962

By C. S. Finney, John Mitchell

The decline of the beehive coking industry was inevitable, but it had filled the needs and economy of its day. A beehive plant required neither large capital investment to construct nor an elaborate and expensive organization to run. The ovens were built near mines from which large quantities of easily-won coking coal of excellent quality could be taken, and handling and preparation costs were thus at a minimum. The beehive process undoubtedly produced fine metallurgical coke, and low yields were considered to be the price that had to be paid for a superior product. Few could have foreseen that the time would come when lack of satisfactory coking coal would force most of the beehive plants in the Connellsville district, for example, to stay idle; and if there were those like Belden who cried out against the enormous waste which was leading to exhaustion of the country&apos;s best coking coals, there were many more to whom conservation was almost the negation of what has since become popularly known as the spirit of free enterprise. As for the recovery of such by-products as tar, light oil, and ammonia compounds, throughout much of the beehive era there was little economic incentive to move away from a tried and trusted carbonization method simply to produce materials for which no great market existed anyway. With the twentieth century came changes that were to bring an end to the predominance of beehive coking. Large new steel-producing corporations were formed whose operations were integrated to include not only the making and marketing of iron or steel but also the mining of coal and ore from their own properties, the quarrying of their own limestone and dolomite, and the production of coke at or near their blast furnaces. As the steel industry expanded so did the geographic center of production move westward. By 1893 it had moved from east-central to western Pennsylvania, and by 1923 was located to the north and center of Ohio. This western movement led, of course, to the utilization of the poorer quality coking coals of Illinois, Indiana and Ohio. These coals could not be carbonized to produce an acceptable metallurgical coke in the beehive oven, but could be so treated in the by-product oven. By World War I the technological and economic limitations of the beehive oven as a coke producer were being widely recognized. After the war the number of beehive ovens in existence dropped steadily to a low of 10,816 in 1938, in which year the industry produced only some 800,000 tons of coke out of a total US production of 32.5 million tons. The demands of the second World War led to the rehabilitation of many ovens which had not been used for years, and in 1941, for the first time since 1929, beehive ovens produced more than 10 pet of the country&apos;s total coke output. Production fell off again after 1945, but the war in Korea made it necessary once more to utilize all available carbonizing capacity so that by 1951 there were 20,458 ovens with an annual coke capacity of 13.9 million tons in existence. Since that time the iron and steel industry has expanded and modernized its by-product coking facilities, and by the end of 1958 only 64 pet of the 8682 beehive ovens still left were capable of being operated. Because beehive ovens are cheap and easy to build and can be closed down and started up with no great damage to brickwork or refractory, it is likely that they will always have a place, albeit a minor one, in the coking industry. The future role of the beehive oven would seem to be precisely that predicted forty years ago by R. S. McBride of the US Geological Survey. Writing with considerable prescience, McBride declared: "A by-product coke-oven plant requires an elaborate organization and a large investment per unit of coke produced per day. Operators of such plants cannot afford to close them down and start them up with every minor change in market conditions. It is not altogether a question whether beehive coke or by-product coke can be produced at a lower price at any particular time. Often by-product coke will be produced and sold at less than cost simply in order to maintain an organization and give some measure of financial return upon the large investment, which would otherwise

Jan 1, 1961

By J. D. Alexander, W. L. Martin, J. N. Dew

This paper presents data obtained using a flood-pot technique to determine the fuel available and the corresponding theoretical air requirements for in situ combustion of crude oils. Since the technique is relatively quick and easy, it is a practical and convenient tool for evaluating reservoirs as fireflood prospects. It is also a research tool which facilitates systematic study of the variables affecting fuel availability and corresponding air requirements. The understanding of these variables is of prime importance to those concerned with the technical and economic development of in situ combustion as an oil-recovery process. The experimental results show conclusively that the fuel available for in situ combustion is not a constant but, rather, varies with crude-oil characteristics, porous-medium type, oil saturation, air flux and time-temperature relationships. Thus, the fuel availability for specified field applications should be determined using actual reservoir crude and core material and the process conditions expected during in situ combustion in the reservoir. INTRODUCTION In situ combustion is a thermal process for recovering crude oil from reservoirs. The thermal energy released during the combustion of a small amount of the oil in place aids in the displacement of the remaining oil. Numerous articles have been published describing the in situ combustion process giving detailed results of laboratory and field experiments.10 In order to engineer an in situ combustion project, a number of important factors must be considered and determined. These factors include: (1) the amount of fuel consumed per unit of reservoir volume swept by the combustion zone, (2) the composition of the fuel consumed, (3) the amount of air required to consume this fuel, (4) the portion of the reservoir swept by the combustion zone, (5) the appropriate air-injection rates and pressures, (6) the amount of oil that will be recovered, (7) the rate of oil production and (8) the operating costs. Nelson and McNiell1 recently have described a procedure which utilizes laboratory combustion-tube data as a basis for the calculation of some of these design factors. Various authors have attempted to describe the in situ combustion process mathematically, and considerable progress has been made. Analytical solutions to the problem of heat transfer from a moving combustion front have been obtained for linear and radial systems."-&apos; All of the published results involve the assumptions that: (1) fuel concentration is constant throughout the reservoir, or that fuel concentration is inversely proportional to the velocity of the front for a given rate of oxygen consumption; and (2) the fuel reacts instantaneously with injected oxygen, while liberating a constant amount of heat per unit weight of fuel at all temperatures. It seems both desirable and reasonable to test the validity of these assumptions experimentally. This paper presents laboratory data which were obtained by means of a "fire flood-pot" method for determining fuel availability and composition, and the corresponding theoretical air requirements for in situ combustion of crude oils under variable conditions. The mechanics of the method are similar to a conventional tube-run experiment.&apos; The important differences involve the size of the reservoir model used and the method for providing the experimental environment. The new method subjects conventionally-sized core samples or unconsolidated sands to a programmed environmental sequence similar to that experienced by a similar volume of rock during the approach and passage of a combustion front in a long tube or in an oil reservoir undergoing in situ combustion. Restored-state samples can also be used. The small samples and relatively simple techniques involved allow an experiment to be set up, run and calculated in about three 8-hour days. This is a considerable improvement over long-combustion-tube techniques which can require several days to run and several more work days to set up and calculate. All the runs presented were run at 40-psig injection pressure. Pressure was not considered as a variable for these experiments, since we previously had found that it had only a small effect on fuel availability up to 600 psig.APPARATUS AND MATERIALS APPARATUS The fire flood-pot apparatus consists of a consolidated

Arizona ROBERT E. TALLY, Chairman CARL E. ~IILLS, Secretary-treasurer, P. 0. Box 1462, Jerome. Ariz. W. S. BOYD THOMAS H. 0. O&apos;BRIEN F. W. MACLENNAN Boston Meets first Monday each winter month GAL~N H. CLEVENG~, Chairman HOLCOMB J. BROWN, Vice-chairman F. I,EROY FOSTER Secretary-treasurer Room 8-219, Mass. lnst: of Tech., Cambridge Mass. A. E. WELLS H. A. WENTWORTH

Jan 1, 1928

By M. L. Brashears

IN the last 15 years industrial use of ground water has more than doubled, and in 1951 amounted to 5 billion gallons per day. A similar sharp increase in the utilization of ground water for irrigation and public-water supply occurred in the same period. In many areas rapid increase in withdrawal from wells has taken place almost entirely unhampered by regulatory control and with little or no integration of effort. As might be expected, the chief interest in many regions has been maximum production rather than sustained perennial yield. As a result, widespread depletion of underground reservoirs and deterioration of the quality of the water stored in them has taken place in many areas, even though total pumpage in the United States is far below ultimate potential. Of even more concern is the fact that excessive withdrawal has drawn salt water into the reservoirs beneath many heavily populated centers along the Atlantic, Gulf, and Pacific coasts, causing costly abandonment of pumping plants.&apos; Many hydrologists expect that consumption of water will rise rapidly in the near future, and some predict that industrial requirements will more than double in the next decade.2,3 Thus it appears likely that the draft on many already heavily pumped underground reservoirs will be greatly increased and the search for additional sources of usable ground water intensified in years to come. In view of this, industry as a whole will be forced more and more to recognize the potentialities and limitations of ground-water reservoirs and to utilize them more effectively to prevent costly water shortages and disruption of production. Through painful experience, some industries are already well aware of the need for effective water utilization, and have managed individually or through joint effort to check trends threatening to deplete underground reservoirs completely or to impair the quality of the water. Various remedial measures have been used to bring about successful management of local or regional ground-water resources. Of these, replen-ishment of aquifers by recharge wells or basins has played an important role in overcoming some ground water problems. Artificial recharge of underground reservoirs by water spreading has been practiced successfully in the United States for many years. In the West it has become an important method of salvaging flood run-off for irrigation of crops and maintenance of public water-supply reserves, and it is used to some extent in parts of the East. Artificial recharge by means of wells, on the other hand, is a relatively new development. Until recently it was employed in only a few areas, principally along the East coast. For the last few years, in the ever increasing search for additional water supplies, industry has had greater recourse to this method. Utilization of recharge wells to control the temperature and quality of underground water supplies is also being considered seriously. Operation of recharge wells, like water spreading, is governed largely by local conditions. It requires water relatively low in turbidity, whereas in some areas water spreading has been used successfully with water of high turbidity and silt content. However, water spreading must be employed in large areas and can be carried on effectively only where aquifers crop out at the surface. Recharge wells can be used in limited space. Recharge wells are similar to production wells except that the water flows in the opposite direction. Thus any water-bearing bed that will yield water to wells may be recharged by wells. Often, however, the water available for recharge is of a different character and temperature from that existing in the ground-water reservoir and if transmitted directly underground from a recharge well to a production well might require expensive or difficult treatment before it could be used. Fortunately the physical characteristics of reservoir beds, which control the movement and behavior of ground water, are generally not homogeneous. Moreover, the movement of ground water is very slow because of the frictional resistance of the reservoir beds. By taking full advantage of hydrologic and geologic conditions, it is therefore possible in many instances to bring about favorable changes of temperature and dilution as the water moves from the recharge wells underground to the production wells. Furthermore, if the natural quality or temperature of ground water is unfavorable for industrial purposes, recharge wells may be used to introduce water of more favorable quality or temperature into the ground-water reservoir. When water is discharged into a recharge well, the head in the well is increased. Because of this, a cone of elevation is produced on the water table or the artesian pressure surface in the area surrounding the well. The cone of elevation is similar to the cone of depression produced around a pumping well except that the apex of the cone is above the water table or artesian pressure surface. Thus if a recharge well and a production well tapping the same water-bearing bed are close together, as would be the case at many industrial plants, some of the water discharged from the recharge well would be drawn into the production well within a short time. Under such conditions it is apparent that water of unfavorable temperature and chemical characteristics should not be used for recharging. The more important ground-water reservoirs in the United States often consist of alternating layers of impermeable beds and porous material that will yield water readily to wells. Physical characteristics of individual beds in a ground-water reservoir may not persist over great distances, the impermeable layers grading into beds that will yield large quantities of water. Thus the water-yielding material in underground reservoirs, whether large or small,

Jan 1, 1954

By P. G. Shewmon

Radioactive MgZA has been used to study the rate of self-diffusion in oriented single crystals of magnesium in the temperature range 468O to 635OC. The diffusion coefficients parallel and perpendicular to the c-axis are: Dl, = 1.0 exp (—32,200IRT) cm2 per sec and Dl = 1.5 exp (—32,500IR T) cm Qer sec. The ratio Dl/Dll was found to vary from 1.13 at 468 to 1.24 at 575 C. Assuming a vacancy mechanism, an explanation of this aniso-tropy of diffusion follows from a consideration of the difference in the saddle points for diffusion in and out of the basal plane. RECENT discovery of radioactive Mg2"&apos;-&apos; has made possible the study of self-diffusion in magnesium. The experimental procedure and the results of a study of diffusion in polycrystalline magnesium have been described in an earlier paper.V he present work is an application of the same techniques to the study of self-diffusion in oriented single crystals of magnesium. In the general diffusion problem the diffusion coefficient is a second order tensor relating the two vectors, the diffusion flux, and the concentration gradient. In a hexagonal lattice, such as magnesium, the complete determination of the diffusion coefficient D as a function of direction in the lattice requires a knowledge of D,, and D,, i.e., the diffusion coefficients parallel and perpendicular to the c-axis, respectively. It can be proven quite generally that in a hexagonal lattice D is independent of direction in the basal plane, and that out of the basal plane it varies with 0, the angle between the c-axis and the direction of diffusion, according to the equation D(B) = D,,cos2B + D,sin2B. [I] The proof of Eq. 1 depends only upon the symmetry elements of the hexagonal lattice and upon the transformation properties of a second order tensor." Therefore, Eq. 1 holds for any mechanism of diffusion and any c/a ratio. Experimental Procedure The experimental procedure used with polycrystalline specimens can be briefly outlined as follows. Radioactive Mg" was produced by bombarding a NaCl crystal with 350 mev protons and was chemically separated from the target as MgO. The Mg" was then vapor deposited on a specially cleaned magnesium specimen by heating the MgO on a tantalum ribbon in a vacuum. During the diffusion treatment, oxidation and vapor loss of the radioactive material were minimized by annealing the samples in pairs with the active faces in contact, each pair being inside a magnesium container, which was in turn surrounded by an argon atmosphere in a sealed Pyrex tube. The distance-activity profiles were obtained by measuring the activity of thin sections cut parallel to the original interface with a lathe. The only technique which was peculiar to this work was the preparation and orientation of the single crystal specimens. The single crystals used in this work were grown by E. C. Burke of the Dow Chemical Co., using a modified Bridgman method, in which the furnace and specimen were stationary while the temperature gradient moved.&apos; In growing these crystals the starting material was distilled magnesium, the crucibles were machined from Acheson electrode graphite, and the furnace atmosphere was tank argon. The crystals were grown from sublimed magnesium with the following analysis: 0.0002 pct Al, 0.0017 pct Fe, 0.0009 pct Mn, 0.0001 pct Ni, 0.0006 pct Pb, and less than 0.01 pct Ca, 0.0001 pct Cu, 0.001 pct Si, 0.001 pct Sn, and 0.02 pct Zn. The two crystals used were roughly lh in. diam and eight in. long. The c-axis made angles of about 7" and 78", respectively, with the specimen axes. If the values of D obtained by the use of these two crystals are taken equal to D,, and D,, respectively, the error introduced by this assumption is less than 1 pct. This can be shown by combining Eq. 1, the identity sin% + cos&apos;8 = 1, and the experimental fact that DJDn in magnesium was always less than 1.25.

Jan 1, 1957