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By L. R. Burdick

The Smokescope is an instrument recently developed for determining the density of smoke by shade comparison. The instrument is designed so that the smoke is seen adjacent to two shades of gray corresponding to Nos. 2 and 3 on the Ringelmann chart. Data on smoke density are important in the smoke-abatement field. The method that is specified and accepted as standard for smoke-density determination by the majority of city smoke regulations involves comparing the smoke shade with the four shades on the Ringelmann chart as published by the Bureau of Mines.2/ This investigation was made to determine the accuracy and convenience of the Smokescope. It was conducted as a part of the Bureau I s air-pollution work, which includes the study of techniques that will promote smoke abatement. The accuracy of the Smokescope was arbitrarily defined as the difference between the Smokescope determinations and the determinations by the accepted, standard method and was expressed as Smokescope correction. Although the corrections obtained were strictly correct only for the two observers and the Smokescope of the investigation, they indicate what might be expected from any observer who uses a Smokescope with reference shades that duplicate those of the test instrument. The report is devoted largely to the data on Smokescope correction, which were taken from a single stack under controlled smoke conditions. Field work, which was done at various locations with different stacks, is mentioned briefly in connection with the comparison of the convenience of the two methods. The data for estimating Smokescope correction were taken for the three variables that were expected to affect the results, namely, the density of the smoke, the amount of light, and the color of the sky behind the smoke.

Jan 1, 1955

By Frank Doxey

AT Dominion Steel & Coal Corp. it has long been recognized that continued mechanization of mine operations is necessary in the Pictou, Cumberland, and Sidney coal fields of Nova Scotia. The varied physical conditions in these fields call for special consideration of individual cases before planning is finalized. Because standard equipment cannot be procured which would operate successfully, many experiments have been necessary over the years to keep pace with the progress made in other countries. There are two mines, producing 2000 tpd, located in the Pictou coal field. The field is badly distorted and crossed by many faults. Seams are highly inclined and irregular and vary in thickness from 5 to 40 ft. Entries are difficult to maintain because of squeezing of the coal. ribs and movement of the roof and pavement. Output from the three operating mines in the Cumberland field is 3000 tpd. The field is highly inclined, inclination varying from 12° to 32°. Overlying beds consist of shales and massive sandstone lenses of extreme toughness and are responsible for bumps when the stresses are relieved by extraction. At greatest depth these are among the deepest coal workings in the world. Depth of cover ranges from 2300 to 4000 ft. This prohibits room-and-pillar working and necessitates longwall operation. Working of contiguous seams concurrently to maintain output increases the already difficult conditions. The Sydney field, with a frontage of about 30 miles, is the most important of the Nova Scotia coal fields. With the exception of one small area it is now wholly submarine. Output is approximately 21,000 tpd. Seams are 21/2 to 8 ft thick, and cover in the areas varies from 600 to 2300 ft, with an average of 1500 ft from sea bottom to the top of the seam. The seams dip in a seaward direction, pitches ranging from 6" to 4.0". The shoreline is the last place of entry to the seams and distance from the bank to the working faces is generally over 3½ miles, in some cases as much as 6½ miles. Ventilation is a problem and requires the construction of large permanent airways. Getting and Loading Coal: In 1925, in view of heavy pressures exerted by thickness of cover overlying the seams, roadways and pillars of the room-and-pillar system being worked began to break up and coal was lost. It was decided that a change in the method of extraction in areas with heavy cover was necessary, and experiments were made with many short walls and longwalls varying between 90 and 250 ft. Trial and error proved that the best operating length was between 400 and 500 ft, delivering all coal to the dip, with roadways to the face following a level course. The change-over was gradual, and the technique of roof control developed with the system, so that falls to the face are now very infrequent. An advantage of longwall mining is that it yields 95 pct extraction, especially important in coal seams of high quality or in seams where faults or disturbances restrict the workable areas. This percentage of extraction is based on the fact that the longwall advancing system takes development faces where pitch permits instead of driving headings and leaving roadway pillars. This system yields high tonnage during development and limits loss in extraction to duff left during operations. Accompanying disadvantages, on the other hand, are the heavy construction cost of main roadways and the necessity of driving all new flank face roadways through the gob. If the main roadways are driven through the solid, and large enough pillars are left on each side for protection against flank face weights, then the width of solid coal is approximately 1700 ft. This represents 10.6 pct of the coal available, or 89.4 pct extraction of the whole. These pillars, however, are of such size that they provide a useful pillar drawing area as a final operation of the mine. It may be that although the seam cannot be generally mined by the room-and-pillar method, it can be adapted to

Jan 1, 1955

By C. S. Chang

Replacement of active groups in corn starch, as shown by results of the starch derivatives tested, impairs rather than improves the value of corn starch as a selective iron oxide depressant. Increased ability to depress iron oxides is generally accompanied by a parallel ability to depress quartz. PEARL starch, a corn starch prepared with a minimum of alteration, is a good selective depressant for iron oxide minerals.' From what is known of the complex chemistry of corn starch, there is evidence that it consists of a mixture of saturated giant polymeric molecules, both straight and branch-chained, the constituents consisting of glucose units with hydroxyl radicals as active or polar groups. To add to the existing knowledge of the chemistry of the flotation process, it is desirable to ascertain the effects on the corn starch as an iron oxide depres-

Jan 1, 1955

By J. G. A. Stevenson

"IntroductionDURING the past quarter of a century diamond drilling, in all its phases, has increased in Canada to a very impressive total of more than 10,000,000 feet annually. Since the art of diamond drilling became recognized, some seventy-five years ago, there has been considerable improvement in both equipment and technique. The steam-powered drill rig and the hand-set diamond bit are things of the past.The purpose of this paper is to review some drilling trends •and practices over the past few years. In the preparation of the paper, the author wishes to acknowledge kind assistance of Mr. J. D. Cumming, Dr. C. H. Hitchcock, Mr. Robert H. Quintal, and many others, including staff members of the Longyear organization.Figures provided by the Dominion Bureau of Statistics for the year 1954 show that Canadian production of metallic and non-metallic minerals and structural materials had a value of $1,454.2 million, excluding radioactive products for which figures are not available. Approximately 61 per cent of this valuation comprised production of nickel, copper, gold, coal, asbestos, zinc, lead, and iron. It is certain that diamond drilling played a major role in the discovery and development of the orebodies from which these minerals were obtained.After the surface examination of an ore deposit, diamond drilling will automatically follow, in order to determine structure, grade, and tonnage.Footage drilled by mining companies and contractors on mineral exploration and blast-hole work since 1945, as reported by the Do-minion Bureau of Statistics, is shown in Figure 1."

Jan 1, 1955

By F. A. Forward, J. Halpern

A new process for extracting uranium from ores with carbonate solutions is described. Leaching is carried out under oxygen pressure to ensure that all the uranium is converted to the soluble hexavalent state. By this method), alkaline leaching can be used successfully to treat a greater variety of ores, including pitchblende ores, than has been possible in the past. The advantages of carbonate leaching over conventional acid leaching processes are enhanced further by a new method which has been developed for recovering uranium from basic leach solutions. This is achieved by reducing the uranium to the tetravalent state with hydrogen in the presence of a suitable catalyst. A high grade uranium oxide product is precipitated directly from the leach solutions. Vanadium oxide also can be precipitated by this method. The chemistry of the leaching and precipitation reactions are discussed, and laboratory results are presented which illustrate the applicability of the process and describe the variables affecting leaching and precipitation rates, recoveries, and reagent consumption. THE extractive metallurgy of uranium is influenced by a number of special considerations which generally do not arise in connection with the treatment of the more common base metal ores. Perhaps foremost among these is the very low uranium content of most of the ores which are encountered today, usually only a few tenths of one percent. A further difficulty is presented by the fact that the uranium often occurs in such a form that it cannot be concentrated efficiently by gravity or flotation methods. In these and other important respects, there is evident some degree of parallelism between the extractive metallurgy of uranium and that of gold and, as in the latter case, it has generally been found that uranium ores can best be treated directly by selective leaching methods. It is readily evident that this parallel does not extend to the chemical properties of the two metals. Unlike gold, which is easily reduced to metallic form, uranium is highly reactive. It tends to occur as oxides, silicates, or salts. Two ores are of predominant importance as commercial sources of this metal: pitchblende which contains uranium as the oxide, U3O51 and carnotite in which the uranium is present as a complex salt with vanadium, K2O-2UCV3V2O5-3H2O. These ores may vary widely in respect to the nature of their gangue constituents. Some are largely siliceous in composition, while others consist mainly of calcite. Sometimes substantial amounts of pyrite or of organic materials are present and these may lead to specific problems in treating the ore. Further complications may be introduced by the presence of other metal values such as gold, copper, cobalt, or vanadium whose re- covery has to be considered along with that of the uranium, or whose separation from uranium presents particular difficulty. In general, there are two main processes for recovering uranium in common use today.'.2 One of these employs an acid solution such as dilute sulphuric acid to extract the uranium from the ore. A suitable oxidizing agent such as MnO, or NaNO, is sometimes added if the uranium in the ore is in a partially reduced state. The uranium dissolves as a uranyl sulphate salt and can be precipitated subsequently by neutralization or other suitable treatment of the solution. The second process employs an alkaline leaching solution, usually containing sodium carbonate. The uranium, which must be in the hexavalent state, is dissolved as a complex uranyl tricarbonate salt, and then is precipitated either by neutralizing the solution with acid or by adding an excess of sodium hydroxide. The latter method has the advantage of permitting the solutions to be recycled, since the carbonate is not destroyed. This is essential if the process is to be economical, particularly with low grade ores. With each of these processes, there are associated a number of advantages and disadvantages and the choice between using acid or carbonate leaching is generally determined by the nature of the ore to be treated. In the past, more ores appear to have been amenable to acid leaching than to carbonate leaching and the former process correspondingly has found wider application. With most ores, acid leaching has been found to operate fairly efficiently and to yield high recoveries. One of the main disadvantages has been that large amounts of impurities, such as iron and aluminum, sometimes are taken into solution along with the uranium. This may give rise to a high reagent consumption and to difficulties in separating a pure uranium product. Excessive reagent consumption in the acid leach process also may result

Jan 1, 1955

By C. H. North, L. E. Lynd, W. W. Anderson, H. Sicurdson

Since the composition and structure of the beach sand concentrates correspond so closely to what would be expected of ilmenites that have been altered by oxidation and leaching, there is no reason to postulate the existence of the hypothetical compound, arizonite, to account for the composition of titaniferous beach sand concentrates. The four concentrates studied are examples of the different products that can result from various degrees of alteration of primary ilmenite. CONSIDERABLE uncertainty is revealed in the literature regarding the nature of the titanium minerals which make up the bulk of the heavy, opaque fractions of numerous beach sand deposits of the world. An investigation was made, therefore, of beach sand concentrates of North Carolina, Florida, Brazil, and India. Magnetic concentration, X-ray, and microscope data were obtained which show that these concentrates consist essentially of ilmenite and its alteration products. The alteration, brought about by oxidation and leaching of iron values, has resulted in upgrading the ores to a range of 56 to 64 pct TiO², as compared with 52.7 pct in ilmenite of theoretical composition. Conclusions drawn regarding the nature of the dominant titanium-bearing mineral in beach sand deposits have been based mainly on chemical composition, specific gravity, and petrography. This material has usually been referred to as ilmenite or weathered ilrnenite.: It has also been referred to as arizonite because in chemical composition it resembles an Occurrence in in which was described by Palmer' in 1909. The chemical composition of Palmer's arizonite corresponded closely to the formula Fe²O³ . 3TiOs. From approximate measurements of crystal angles on a large imperfect crystal with rounded faces, the

Jan 1, 1955

By P. E. Busby, C. Wells

FURTHER study of data used in determinations of 1—rates of diffusion of boron in austenite and 2—solubilities of boron in the a and phases of iron and steel' has provided an equation for the diffusion of boron in a iron. In brief, the previously published data were accumulated from the results of deboron-izing (and decarburizing) experiments carried out in the range of 700° to 1300°C. Diffusion coefficients (D?) for boron in austenite were calculated using the Grube solution of Fick's law. However, only solubility values were estimated from the discontinuous concentration-penetration curves, Fig. 1, which are characteristic of diffusion in two phases. Dr. Carl Wagner' has recently provided a solution for calculating D values from penetration curves of this type as follows: Cl1.1 —C8/C8- C11.1 = vp ? e ? erf (?) and D = ?/4?2 t where C11,12, C8, C11, and E have the values shown in Fig. 1. D is the diffusion coefficient, sq cm per sec; t is time of deboronizing anneal, see; and ? is a di-mensionless parameter. For a given diffusion experiment, the value of ? can be readily obtained by graphical solution from a plot of y vs the right-hand part of Eq. 1. D may then be evaluated from Eq. 2. The avplication of this solution to previously re-ported results, of which excerpts are given in Table I, permits the calculation of diffusion coefficients for boron in a iron. On the basis of these meager data, it is tentatively concluded that the diffusion of boron in a may be represented by the equation Da = 1011 e-92,900/RT where R is the gas constant, cal X 0C-1 X mol-1; and T is the absolute temperature, OK. Although the frequency factor, 106 sq cm per sec, is admittedly several orders of magnitude higher than expected, the value of Q, 62,000 cal per mol, appears reasonable and is, in fact, very similar to that for the self-diffusion of iron. It is pertinent to mention at this point that the value of Q obtained for the diffusion of boron in austenite by means of the Wagner solution is 19,000 cal per mol and is in excellent agreement with the value previously reported1 in the equation DT = 2xl0-3 e-21,000/RT [4] which was determined by the application of the Grube solution to other data. The fact that determined constants A and Q in the equation D? = A e-Q/RT were practically the same, independent of whether the Wagner or Grube solutions were used in determinations of D values, strengthened the authors' belief that the computed values of D (Table I) using the Wagner solution are reliable. The relative values in Eqs. 3 and 4 for the diffusion of boron in the a and ? phases, respectively, suggest that boron forms a substitutional solid solution in a iron and an interstitial solid solution in ? iron. The same tentative conclusion has been reached by McBride et al.3 on the basis of relative solubilities, atom diameter, and the size of the interstitial hole in a and ? iron. In connection with the data for test 14 listed in Table I, it is of interest to calculate the solubility of boron in a iron using the D value given by Eq. 3. As might be anticipated from the small movement of the interface in test 14, proper substitutions in Eqs. 2 and 1 give a low value, approximately 0.0004 pct B, at 850°C. Apparently at 835°C (test 13) it is possible to obtain 0.0018 pct B, and at 850°C only 0.0004 pct B into solution in the a phase before a second phase appears. These observations are consistent with the Fe-B constitution diagram proposed by McBride, Spretnak, and Speiser.3 References P. E. Busby, M. E. Warga, and C. Wells: Diffusion and Solubility of Boron in Iron and Steel. Trans. AIME (1953) 197, p. 1463; Journal of Metals (November 1953). 'W. Jost: Diffusion in Solids, Liquids, Gases. (1952) pp. 69-71. New York. Academic Press Inc. "C. C. McBride, J. W. Spretnak, and R. Speiser: A Study of the Fe-Fe2B System. Trans. ASM (1954) 46, p. 499.

Jan 1, 1955

The Black Hills Mineral Atlas briefly summarizes all important data available on the mines and mineral deposits in the Black Hills that have been explored, developed, mined, or located. The purpose of the atlas is threefold; First, from available data the general potentialities of the Black Hills region can be determined, although it is not possible to calculate actual ore reserves; second, it will provide the necessary basic information regarding a given property or commodity for further detailed study; and third, the compilation is a comprehensive directory of Black Hills deposits.

Jan 1, 1955

By H. Wilton Clark

VARYING in thickness and in number from place to place, coal seams in the Canadian Rockies also range in pitch from nearly horizontal to vertical, sometimes with overturns. Over the entire coal-bearing area there are considerable differences of rank in coals of the same geological age, and there are marked differences in ash content and wash-ability characteristics. Correlation of seams at mining operations within a few miles of each other has often proved impossible. These factors influence mining methods, and, of course, production results. The oldest coal formations of western Canada are of Lower Cretaceous age, as the carboniferous sediments are marine and contain no coal. Coal is also present in formations of Upper Cretaceous and Tertiary ages. The rank of coal varies from semi-anthracite in some operations in the Kootenay and Luscar formations of the Lower Cretaceous in the Rockies to lignite in the Tertiary fields of the Saskatchewan prairies. Fig. 1 shows the general extent of the formations. The pitching seams, chiefly Lower Cretaceous, occur in the western belt of the Rocky Mountain foothills and the eastern slopes of the Rockies themselves. (The road to Tent Mt. strip pit, elevation 7000 ft, is shown on p. 832.) Formations extend from the U. S.-Canadian boundary for several bundred miles in Alberta and continue for a similar distance in British Columbia. Present geological estimates show a probable reserve of the order of 25,000 million tons of coal ranging from high volatile bituminous to semi-anthracite, with a large percentage of coking coal, constituting one of the major world reserves of that type. It should be noted that although the quoted estimate probably errs on the conservative side, the question of access to the seams in a mountainous terrain will always be a problem; a wide divergence may exist between actual reserves and the quantities of coal economically recoverable. An excerpt from the Canadian government Report of the Royal Commission on Coal 1946 states that conditions favorable to coal formation were intermittent and these intervals relatively brief in comparison with periods in which no coal was formed. Conditions were favorable to the growth or accumulation of vegetation in one area, while fresh water sediments and marine shales were being deposited in other areas. During periods of emergence the coal deposits were subjected to erosion or were covered by coarse sands and gravels from the mountains, whereas during submergence they were covered by fine clays deposited in embayments of the sea. During some of the periods of coal formation, volcanic activity deposited beds of fine volcanic ash and dust with the coal-forming vegetation. Coal deposits reach their greatest development in the mountains and thin rapidly to the east into the plains area, where they are deeply buried beneath younger sediments. For example, in the Fernie area of southeastern British Columbia there are in places 22 seams having an aggregate thickness of 150 ft in a stratigraphic interval of 3500 ft, whereas at Cole-man in the Alberta Crowsnest area the measures are only 800 ft in thickness and contain a maximum of five seams aggregating about 47 ft of coal. At Belle-vue, 10 miles further east, the measures are reduced to 430 ft with only three seams aggregating about 37 ft of coal. Subsequently the tremendous forces involved in the upthrust of the Rocky Mountains produced great displacement of the coal-bearing formations and at the same time a change in geologically young coals from low to high rank. The floor or footwall of most seams consists of carbonaceous shales, which are almost as adaptable to bending as the coal seams themselves. The seams show very few cleats or cleavage planes and have been so weakened structurally that as mined 50 pct or more will pass through a ¼-in. screen, seldom

Jan 1, 1955

By L. E. Wilsey, W. G. Bearden

This paper presents results of analysis of the effect of fracturing on initial flow rates and on ultimate recoveries from low capacity oil formations. This analysis shows that even in formations of permeability as low as 0.1 md, large fracture treatments can yield oil recoveries and production rates approaching those of high permeability formations. Field production information is shown which supports the analytical work. INTRODUCTION Often hydrocarbon bearing formations are discovered which, while they are known to contain large quantities of oil, will not produce at commercially attractive rates without special well stimulation. Attempts to solve this problem led to the development of such techniques as underreaming, shooting, acidizing, and recently hydraulic fracturing. These methods of stimulation have been highly successful in many fields. All of these treatments, however, affect only the pay section near the well bore and therefore owe their success to circumventing and overcoming local permeability reduction about the well bore. It has been the belief that in treating wells to stimulate production in which there is no permeability reduction near the well, the maximum possible production rate increase would be of the order of two or fourfold. Hence, it has been concluded that fracture treatment of uniformly low capacity (permeability times thickness) formations could meet with but limited success. It is the purpose of this paper to show that the use of fracture treatments much larger than usually used can yield commercially attractive producing rates and high recoveries even in formations of permeability less than 0.1 of a md. PROCEDURE In order to illustrate the effects of fracturing, an analytical study of some hypothetical systems was made in which one factor at a time was varied to show what is important in recovering oil from fractured wells. It was assumed that a horizontal fracture of relatively high fluid-carrying capacity extended from a well bore. The presence of such a fracture will cause a marked influence on the flow pattern and pressure distribution in the formations, with the net effect being to reduce tremendously the resistance to flow into the well bore. A diagram of the resulting flow pattern and pressure distribution is illustrated in Fig. 1. Immediately after production commences, the principal flow of fluid is vertically from the formation flanking the fracture, into the fracture, and then into the well. After sufficient oil and gas have been produced to lower the pressure in the formation adjacent the fracture, oil and gas will begin to flow in a radial manner into the fracture from the formations beyond the end of the fracture. In calculating the performance under such conditions. the material balance or conservation of matter equation and equations of flow were solved simultaneously by a stepwise method of calculation. Flow rates were found by determining the pressure distribution, as illustrated

Jan 1, 1955

By W. F. Kieschnick, and W. J. McGuire&apos, Eugene Harrison

This paper concerns the induction and extension of fractures into rock formations as involved in drilling, completing, and production stimulating operations on wells. Conclusions concerning formation breakdown are derived from (I) a review and extension of published analyses relating to mechanical theories of rock stress and the state of stress in the earth's crust and (2) a correlation of field data from fracturing operations. Conclusions concerning the mechanics of fracture extension, which indicate the relationship between fracture dimensions and rock properties, depth, and volume of injected fluid, are tentative and largely establish limits of relationships. These conclusions are derived from stress calculations, limited field data, and laboratory experimental studies. The experimental work involves the study of the stresses at the fracture boundaries and the geometry of pressurized fractures by means of photo-elastic modeling methods. Results of this investigation indicate that a large majority of pressure induced wellbore fractures are vertical, particularly in deeper wells; and variations in the pressures necessary to create and extend fractures can be explained largely on a basis of established rock properties. It is also shown that variations due to tectonic forces should usually be expected to be slight. Other results indicate that during that extension of fractures rather large fracture volumes are temporarily created by the parting of the formation. The purpose of this paper is to present the results of calculations and laboratory experiments concerning the mechanics of fracture induction and extension with a view to broadening existing knowledge relating to these phenomena. It is believed that continued progress in developing knowledge of this type is important to the further development of techniques for drilling and completing wells. When a wellbore is subjected to sufficiently high pressure, fracturing of the formation surrounding the well occurs. This phenomenon is a common cause of lost circulation during drilling operations, is a deliberately planned factor in most squeeze cementing operations, and is the controlling factor in many productivity stimulation operations. The exploitation of formation fracturing in recent years has been particularly impressive in sand-viscous fluid well stimulation treatments which were first introduced to the industry by the Stanolind research group1,2,3 several years ago. The borehole fracturing phenomenon has been analyzed mathematically by application of the principles defining the elastic and inelastic behavior of thick-walled cylinders of homogeneous, isotropic, and impermeable material.1,5,6 This basic argument is re-presented in this paper along with new considerations which strengthen it. In addition to these arguments relating to fracture induction some calculations and experimental results relating to the mechanics of fracture extension are presented. Since the nature of the pressure induced fracture at distances from the well is of great importance in considerations of well stimulation procedures there appears to be a need for an integrated approach which deals with the mechanics of fracture extension as well as with wellbore rupture.

Jan 1, 1955

By S. F. Reiter, R. W. Guard, J. H. Keeler

DURING the last several years the authors have been conducting tests on a tensile machine that autographically records load-elongation data. In certain tests at elevated temperatures on face-centered cubic, hexagonal close-packed, and body-centered cubic materials (binary alloys of nickel, zirconium, and iron), the load-elongation curve as shown in Fig. 1 reaches a maximum load (Pmax,) at relatively small amounts of strain and a large amount of elongation occurs subsequently with a continuously decreasing load, until finally localized necking and fracture take place. This early maximum load behavior has been found without the accompanying evidence of localized necking as has been assumed by some, or of evidence of cracking as observed by others.' This behavior has also been observed by Nadai and Manjoinez whose results are in good agreement with ours. The prior strain, strain rate, test temperature, and crystal structure affect the observations. The influence of prior strain was observed in tests on unalloyed zirconium at 300°C in which specimens containing large amounts of prior strain (cold work) exhibited the type of curve shown in Fig. 1. No localized necking was observed after strain as much as that of point B. Annealed and recrystallized specimens of the same material showed a load-elongation curve with maximum load occurring after considerable strain hardening and at the beginning of localized necking. Interrupted tests (Fig. 2) show no loss in strength because of recovery or recrys-tallization during an interval at temperature without load. Recovery under stress such as found by Wood and Suiter3 in aluminum is now being investigated as one possible explanation. The strain rate is a contributing variable as shown in Fig. 3, a test on high purity nickel. Here the strain rate was decreased by a factor of ten during a test in which the maximum load was reached at a small strain. The slope of the true stress-true strain curve was positive with the more rapid strain rate and negative with the slower strain rate. In many binary ferrites an added effect of temperature has been observed. Fig. 4 shows the load-elongation plots for a binary ferritic alloy containing 2 pct Mo. The slope of the small portion of the stress-strain curve obtained at a strain rate of 0.01 mhr1 changes sign from negative to positive on going

Jan 1, 1955

By A. R. Troiano, R. F. Hehemann

The influence of partial decomposition to high temperature bainite on reaction kinetics at a lower temperature has been studied in two alloy steels. Reaction at the lower temperature is retarded by the prior treatment, and the extent of decomposition may be reduced. Interpretation of these results is based on a mechanism involving a limitation in the nucleation and growth of bainite plates. OF the major transformations in steel, the characteristics and general behavior of the bainite reaction are probably the least understood and appreciated. Limitations of space preclude a critical evaluation of the present status of the bainite transformation in this presentation; however, such a treatment will shortly appear elsewhere. Only the salient features pertinent to the present investigation will be introduced briefly here. Although the reaction curve for the formation of bainite is similar to that for a nucleation and growth process, other kinetic features are more in keeping with the martensitic mode of transformation. A definite temperature exists above which austenite will not transform to bainite.1-5 his temperature, which has been designated B., is determined by the composition of the austenite. Unlike other nucleation and growth processes, the amount of austenite transformed to .bainite is a function of reaction temperature. The extent of decomposition increases from 0 at H. to 100 pct at some lower temperature.' , This lower temperature will be designated B1 and appears to be relatively insensitive to austenite composition.% 5 The similarity in the effect of reaction temperature on the bainite and martensite transformations serves to emphasize the close connection between these two decomposition processes. Austenite decomposition in the bainite range proceeds without partition of the alloying elements.8-11 Partition of carbon has been proposed" primarily on the basis that partial transformation to bainite lowers M, and increases the amount of austenite retained at room temperature. Carbon enrichment resuslting from such partition has been employed to explain the influence of reaction temperature on the extent of decomposition.'" It should be noted, however, that no enrichment has been detected experimentally in high carbon steels.1,14,15 Lattice-parameter measurements of retained austenite in steels containing 0.3 to 0.4 pct C have indicated carbon enrichment, 3,10-18 although the split indicative of a high carbon martensite has not been reported. Carbon enrichment, if it does occur, must be highly localized around the bainite plates. Therefore, carbon enrichment does not account for the influence of temperature on the progress of the bainite reaction."' Thermal history is known to influence the martensite transformation through stabilization.20,21 No similar phenomenon in the bainite transformation has been reported. Materials and Procedure Two triple-alloy steels were chosen for this investigation. Their compositions were as given in Table I. These steels were chosen because the pearlite reaction did not interfere with the bainite reaction. Steel K was received in the cast condition and forged from 2 in. square bars to 1/2 x1 3/4 in. plates. The 4340 was received as 11/4 in. hot-rolled rounds. Both steels were homogenized in vacuum for one week at 2300°F in order to minimize segregation. A quenching dilatometer similar to that described by Flinn, Cook, and Fellows" was employed for the kinetic measurements. Dimensional changes were detected by a differential transformer coupled with a high speed recorder. The dilatometer was mounted so that it could be transferred to any one of three furnaces: a nitrogen-atmosphere austenitiz-ing furnace and two salt-bath furnaces for isothermal transformation. Dilatometer specimens were 1/32 x 1/4x 1/2 in. with a gage length of 1.4 in. All specimens were nickel plated in order to minimize decarburization during austenitizing. The austenitiz-ing conditions consisted of 10 min at the temperatures given above. Austenitizing temperatures were controlled to 210°F and transformation temperatures to ±3ºF. The precision of the dimensional measurements was estimated to be ± 5 x105 in. per in. Results and Discussion Isothermal Transformation: The characteristics of the isothermal bainite reaction will be described

Jan 1, 1955

By H. P. Holmes

Tonnage quantities of hafnium metal have been prepared in Kroll-process equipment designed for producing zirconium. Some slight process modifications have been necessary, but overall operation is the same as for zirconium. Hafnium oxide is chlorinated in the presence of carbon and the resultant hafnium tetrachloride reduced by magnesium metal. Byproduct magnesium chloride is removed in a high-vacuum distillation step and the hafnium sponge metal crushed for further use. Arc-melted metal can be forged and rolled to sheet, but the best material produced to date would not be considered cold ductile. INTRODUCTION This investigation is part of the program for the production of pure metals from domestic and foreign ores that has been underway at the Bureau of Mines for several years. It was done in cooperation with the Atomic Energy Commission and the Bureau of Ships.

Jan 1, 1955

By C. S. Matthews, P. Hazebroek, F. Brons

A method has been developed for calculating the average pressure in a bounded reservoir. The reservoir is first divided into the individual drainage volumes of each well, by using the criterion that at steady state each individual drainage volume is proportional to a well's production rate. The average pressure in each drainage volume is then calculated by a method developed in the report. By volumetrically averaging these individual drainage volume pressures, the average pressure in the entire reservoir is obtained. To calculate the average pressure in each drainage volume, a correction is applied to the ordinary extrapolated pressure, i.e., the pressure obtained by extrapolating to infinite time the linear portion of the graph of closed-in pressure versus log[?t/(t + ?t)], where At is the closed-in time and t the production time. The correction, which is a function of the production time, is presented in graphical form for different shapes of the drainage area (horizontal cross section of the drainage volume). INTRODUCTION It is important to be able to find the volumetric average pressure in a reservoir so that the size of the reservoir may be determined from material balance calculations. It is also desirable to be able to find the approximate distribution of pressure within a reservoir for detection of fluid movement. 'The purpose of this paper is to present a method for calculating both the average reservoir pressure and the approximate distribution of pressure within a bounded reservoir — that is, a reservoir with no water drive. In reservoirs where the pressure builds up rapidly after wells are shut in, the determination of average pressure generally poses little problem, for one often need only average the final buildup pressures. It is when pressure buildup is slow that difficulties arise. For practical and economical reasons, the time allowable for closing in wells is limited. If at the maximum allowable closed-in time the pressure has not reached a con- stant value (and this is more often the case than is generally realized), calculation of average pressure presents difficulties. One method which has been used in the past for obtaining an "average" pressure involves the extrapolation to infinite closed-in time of curves of pressure as a function of closed-in time. Contour maps of these extrapolated pressures are used in arriving at an average pressure. In some cases wells are shut in for arbitrary Lengths of time, and contours of these arbitrary shut-in pressures are used for calculation. All these methods are somewhat arbitrary and suffer from lack of fundamental theoretical basis. For example, no precise meaning can be given either to the extrapolated or to the arbitrary shut-in pressure. Furthermore, there seems to be no reason why either set of pressures should be contourable. It was because of the vagueness of these old methods and their lack of a sound theoretical basis that, the present method for determining average reservoir pressure was developed. SYMBOLS A = drainage area of a well, cm' A, = drainage area of well j, cm2 A , = total drainage area of all wells in a reservoir, cm2 c = coefficient of compressibility of fluid at reservoir temperature, atm-1 f = hydrocarbin-filled porosity, fraction h = net thickness of formation, cm k = permeability of formation, darcy p = pressure at a well, atm p0 = initial reservoir pressure, atm p* = pressure obtained by straight-line extrapolation of linear portion of plot of p versus log [At/(t+ At)] p = volumetric average pressure inside drainage volume of a well q = volume flow rate at a well, cc/sec a: prevailing reservoir conditions q, = total volumetric flow rate from a reservoir, cc/sec at prevailing reservoir conditions rw = radius of well bore, cm T = kt/fµcA t = corrected time of production of a well, seconds, defined by

Jan 1, 1955

By T. T. Campbell

A survey of the literature eras made covering the use of aqueous non-aqueous, and fused-welt systems for the electrowinning or electroplating of molybdenum metal. The survey was prepared in order to provide background for further investigations directed toward the electrodeposition of pure molybdenum metal from either aqueous or nonaqueous mediums. While a thorough effort was made to cover all available references, a critical review was not attempted nor does this survey pretend to be exhaustive in nature.

Jan 1, 1955

By J. L. Eakin

Systematic flooding was introduced into Nowata County, Okla., in May 1931. The results of this means of stimulating the nearly depleted Bartlesville reservoir in the county has been the subject of four Bureau of Mines Reports of Investigations.2-5/ This report presents a discussion of four full-scale water-flooding projects in the Delaware-Childers field and 1 pilot project in the Curl Creek field (fig. 1). All five plants have been in operation for only a few months. Although substantial increases in the rates of oil production have been obtained 1 by the injection of water, the projects are of particular interest at this time because of the novel methods of operation and new equipment that are being employed. One project utilizes a portable "package" plant designed for a pilot operation. All of the controls for this project are centrally located. The design makes it possible to remove the test plant to another location if the project is unsuccessful or if the unit is to be replaced by a larger plant. This project also utilizes diatomaceous-earth filters successfully for the first time and makes use of plastic pipe. Another novel feature of the operation is the aeration of Arbuckle limestone water through an airlift in the supply well.

Jan 1, 1955

By E. E. Harton

Bureau of Mines Information Circular 16, Bibliography of Investment and Operating Costs for Chemical and Petroleum Plants, was published in October 1949 as an aid in making estimates of investment and operating costs of processes for manufacturing synthetic liquid fuels. Development of such processes by the Bureau of Mines was authorized by Public Law 290, 78th Congress. This bibliography included refer¬ences to and abstracts of publications appearing from 1930 to September 1948 on all processes in general use in the petroleum refining and chemical industries. Publications in this field have been numerous since compilation of the above bibliography, and the present report was prepared to cover the field from September 1948 through June 1952.3/

Jan 1, 1955

By G. A. Pinsky, Pickard M. K

"IntroductionTWO METHODS of mucking have been used in sinking C shaft, the five-compartment production shaft of the Yellowknife property of Giant Yellowknife Gold Mines, Limited:(a) Band mucking during the original sinking in 1949 and 1950 (l); and(b) Mechanical mucking with the Cryderman shaft mucker during the recent deepening in 1954.This paper briefly describes the recent deepening and compares the performance figures of the two sinking operations.PlanningProduction planning in 1952 showed that it would be advisable to deepen C shaft in 1954 to provide for three new production levels and the installation of an underground crusher. To complete the required shaft work it would be necessary for a shaft crew to deepen the shaft about 500 feet, to partly excavate three level stations and a crusher station, and to cut a station for the loading pocket. Figure 1, a vertical section of the lower part of the shaft, shows the details of the deepening programme.During the planning period, Giant had a rather small supervisory staff and at the same time was experiencing a heavy labour turnover. As it was obvious that production would suffer if the sinking crew were recruited at the mine, it was decided that the merits of contracting-the shaft sinking to an independent contractor should be investigated."

Jan 1, 1955

By Elmer R. Kaiser

G. A. Vissac (Consulting Engineer, Vancouver, B. C.) —Some of the data presented in this paper, in connection with cost studies of washed coals, should be clarified and qualified. Washing a raw coal down to a low ash content may mean higher losses in refuse, but the cost of the operation may be more than compensated by the increased value of the resultant product. However, careful engineering and accurate machinery are required to insure maximum profits. Washing at a low gravity separation demands a very accurate automatic process, chiefly when the small sizes of coal are being handled. Coal consumers today are buying mostly on a Btu basis, cognizant of the fact that low ash contents result in lower installation costs and in savings in handling and transportation charges. The attached Table X gives the actual balance sheets of the various cleaning operations corresponding to various ash contents for the two coals presented in the paper, using the old technique of a two-products separation. or the new technique, now commonly used in Europe, of a three-way separation made possible by new methods and new equipment. Savings in capital costs, handling, and transportation have been calculated on the basis of a delivered cost f.O.b. plant of $10.00 per ton of coal at 16.5 pct ash (actual 1953 average price reported by Consolidated Edison, N. Y.), and of $3.50 at the mine also for 16.5 pct ash coal. In both cases this corresponds to values of 126 and 4.26 per 1 pct of combustible and per ton, respectively, on delivered and on at the mine prices. As an example, the middle western coal at 16.5 pct ash, raw, washed down to 4 pct ash in a two-way process will give 53.75 pct of clean coal at 4 pct ash and 46.25 pct of refuse at 31 pct ash. The value of the losses will be 0.4625x(100—31)x0.042 = $1.34. Add the cost of washing, capital, and operation, which is $0.35. The total cost will then be $1.69. But with a three-ways separation we will have 53.75 pct of first-grade coal at 4 pct ash, 37.25 pct second-

Jan 1, 1955