Search Documents

By C. Wagner

When an alloy solidifies and the composition of the solid differs from the composition of the liquid, atoms of the alloying elements rejected at the solid-liquid interface have to diffuse toward the bulk liquid. Diffusion may be supported by convection. Theoretical calculations have been made for a liquid without convection, for a liquid involving natural convection, and for solidification of a liquid alloy at the surface of a rotating disk as an example of forced convection. WHEN an alloy solidifies, the equilibrium composition of the originating solid phase differs in general from the composition of the liquid phase. Thus, segregation may take place, i.e., the composition of the solid phase formed at different times differs from the bulk composition. The degree of segregation depends on several factors such as the partition ratio of the alloying elements between the liquid and the solid phase, the rate of solidification, and convection. To clarify, a theoretical analysis has been worked out. The following presuppositions are made: 1—Calculations are confined to the solidification of binary alloys in which the concentration of component A (solvent) is much greater than that of component B (solute). 2—The equilibrium concentration of the alloying element B is assumed to be lower in the solid phase than in the coexisting liquid phase. 3—At the solid-liquid interface, partition equilibrium is assumed. 4—The dependence of the partition ratio of the solute between the liquid and the solid phase on concentration and temperature is disregarded since only small concentrations of the solute are considered. 5—The concentration of the alloying element B in the solid phase is supposed to be lower than the saturation concentration for the formation of a second solid phase. 6—diffusion in the solid phase is disregarded, since diffusion rates in the solid state are comparatively low. When the partition ratio of the solute between the liquid and the solid phase is greater than unity, a part of the solute is rejected at the solid-liquid inter- face and has to diffuse toward the bulk liquid. Thus there is necessarily a concentration difference in the liquid between the solid-liquid interface and the bulk liquid as has been pointed out by Hayes and Chipman,' Rutter and Chalmers,' Tiller, Jackson, Rutter, and Chalmers, and others (see Fig. 1). In view of the concentration gradient in the liquid phase, the ratio of the concentration of the solute in the solid phase to that in the bulk liquid differs from the equilibrium ratio, although at the solid-liquid interface, partition equilibrium is assumed. Under certain limiting conditions, the solid may even have virtually the composition of the bulk liquid as is shown below. General Equations Let u be the linear crystallization velocity in cm per sec, D the diffusion coefficient of component B in the liquid phase in cm2 sec-', c',, and c", the concentrations of component B in the liquid and the solid phase, respectively, in mol per cm", and s the distance from the solid-liquid interface toward the bulk liquid at a given time t. The partition ratio is assumed to be equal to a partition constant K. Thus. The difference in the specific volume between the liquid and the solid phase may be disregarded. In view of the discontinuity of the concentration of the solute at the interface according to Eq. 1, u(c', — c",), , , = tic', (S = 0)(1 — K) gram-atoms B per unit area per unit time are rejected at the interface and have to diffuse toward the bulk liquid. From Fick's first law it follows that uc',,(s = 0) (1 ~K) == - D(dc',,/?>s). _-„. [2] The concentration profile in the liquid near the interface depends on the interplay of diffusion and convection as is discussed in the following sections. A schematic concentration profile is shown in Fig. 1. The "effective thickness of the diffusion boundary

Jan 1, 1955

By Bension Varon

Systematic projections of mineral supply, demand and prices are an integral part of the mineral sector planning process. As such, their primary value is not as prophecies but as devices for imposing discipline on thinking about the future and as means of ensuring consistency in policy decisions. Both the present and the broad outline of the future as presently discernable affect projections on specific minerals. The present world mineral situation reflects most prominantly the recent commodity cycle while our view of the future is largely shaped by perceptions and attitudes of policy-makers and the public over larger issues. Mineral commodities are of considerable importance to various countries. This in itself dictates a strong interest in systematic projection activities, the need for which is heightened by the volatility of certain basic variables affecting mineral demand in the world. It is useful for presentation purposes to discuss the various techniques suited for projecting mineral trends under three principal headings, viz., econometric modeling, systems analysis and technological forecasting. In addition, two specialized variants, input-output analysis and resource base modeling, are discussed. In each case, a brief description of the approach and some indication of its application to projections of mineral commodities are provided.

Jan 1, 1977

By Ward W. Minkler

Since 1978, the titanium metals industry, throughout the non-socialist world, has experienced a period of growth and prosperity unparalleled since the late 1950s. Sponge producers, melters and mill product manufacturers have operated at capacity, record shipments have been attained, new producers have entered the industry and established producers have implemented expansion plans. However, there is now mounting evidence that the underlying elements which generated this vigorous surge have abated, at least temporarily. It now appears that in 1982, the industry will encounter overcapacity, excessive inventory accumulations and the probability of declining prices. Therefore, it does seem appropriate that we now take inventory of the present and future markets for titanium with the hope we can identify both near term opportunities as well as challenges the industry now faces. If we view the development of titanium markets as a function of time, they fit rather neatly into three sectors.

Jan 1, 1982

By J. Menashi, D. A. Douglas, A. S. Rappas

Aqueous solutions of Na2Cr04 are reduced with CO to produce hydrous Cr2O3 and NaHCO3/Na2CO3. CH3OH, HCOO- and HCHO in the presence of CO2 also reduce Cr(VI) and yield similar products. It is shown that the rate of Cr(VI) reduction can be represented by the expression -d [Cr(VI) ]/dt = k[Cr(VI)]0.5[R], where [R] is the aqueous phase concentration of CH3OH, HCOO- or CO. With these three reductants, the activation energies for the reduction processes are comparable and are of the order of 100 kJ/mol (43,000 Btu per lb mol). Rapid reduction rates are attained at temperatures in excess of 220°C (428°F). With these three reductants, a green acid-insoluble readily filterable hydrous Cr203, is formed. With HCHO, significant reduction of Cr(VI) occurs at about 80° to 100° C (176° to 212° F). The product is an acid-soluble violet modification of hydrous Cr203. Solutions of Na2CrO4 contaminated with Mo(VI) and W(VI) can be reduced to yield a Cr2O3 product with significantly reduced contaminant levels.

Jan 1, 1986

By M. J. Salkind

In discussing the possible contribution of plastic deformation to neck growth during sintering, Wilson and Shewmon State that Lenel, Ansell, Salkind, and Early34, 35 concluded that dislocation flow is the rate-controlling process in sintering under zero external load. The experiments cited34, 35 were done with applied external loads, and the conclusions were that the same kinetics could be controlling sintering due to the force of gravity with no additional applied load. The effect of gravity on sintering behavior was previously cited.34, 31 In addition, Wilson and Shewmon argue that the conclusions of Refs. 34 and 35 are doubtful because the strain rate is proportional to stress for plastic flow. This assumption is true only for certain kinds of plastic flow such as Nabarro-Herring Microcree38 As Early, Lenel, and Ansell point out3' plastic flow, based upon dislocation movement, can exhibit strain rates which exhibit a power law relationship to the applied stress.39, 40 Wilson and Shewmon also doubt the results of Refs. 34 and 35 because the stresses in the necks are too small for x/a > 0.1. This is true, and the experiments described in Refs. 34 and 35 were done for small values of %/a. The conclusion was that plastic flow by dislocation movement was important during the early stages of sintering when x/a is small.

Jan 1, 1968

By A. S. Yue, A. E. Vidoz, F. W. Crossman

Tensile specimens were prepared from a single grain of an epitaxially grown Al-CuAl2 eutectic ingot. The eutectic lanzellae were oriented parallel and perpendicular to the tensile axis of the specimens. Since the composite was of the eutectic composition, the aluminum-rich matrix could dissolve up lo 5. 7 wt pct Cu in solid solution and, therefore, was amenable to strengthening by precipitation hardening. The tensile properties of the eutectic single crystals were determined at room temperature as functions of interlamel-lar spacing, platelet orientation, and thermornechanical trealment. The obserced variations in composite stress and modulus with respect to the level of&apos; composite strain are discussed in terms of premature fracture of CuAlz platelets, a distribution function for the strength of the lamellae, and unequal strains due to localized fracture of&apos; platelets. The discontinuous fiber composite model of Kelly and Tyson is modgied to account for a changing distribution of fiber lengths during composite loading. The tensile properties at elevated temperatures were determined for the direc-tionally solidified eutectic oriented with platelets parallel to the tensile axis. The observed properties are attributed to the onset of plasticity of the CuAL2 phase above 150°C. DURING the investigation of whisker- and fiber-reinforced metallic matrix composites in recent years, two major problem areas have developed: 1) The fabrication of the composite involves tedious handling techniques in order to obtain a unidirection-ally aligned and uniformly spaced set of whiskers in the metal matrix. 2) Due to weak interfacial bond strengths or because of the formation of additional embrittling phases at the metal-fiber interface during long-time exposure or fabrication at elevated temperatures, many composite systems have exhibited considerably lower strengths than those predicted by a law of mixtures analysis.&apos; These problems have been bypassed by the technique of growing whiskers and plates of high-strength materials in a ductile metal matrix by controlled unidirectional eutectic solidification.2 The tensile properties of directionally solidified A1-CuA12 eutectic are presented here. This alloy consists of a ductile aluminum matrix, containing up to 5.7 wt pct Cu in solid solution, which is amenable to precipitation hardening by heat treatment and a reinforcing high modulus CuAlz intermetallic phase. The two phases are present in the form of alternating platelets or lamellae. The microstructural stability of this unidirectionally solidified alloy at elevated temperatures has been studied extensively.3.4 and preliminary tensile and bend tests have been reported. 5-7In the present investigation the tensile properties of the A1-CuA1, eutectic have been studied as a function of several ther-momechanical variables: solidification rate. heat treatment. rolling at elevated temperatures. and lamellar orientation. It was felt that the uniformity of structure and excellent interfacial bonding would give tensile properties concomitant with the metal matrix composite theory of strengthening proposed by Kelly and coson. 8-9 The tensile properties that were obtained point to a wide distribution of strengths for the CuAlZ platelets, which leads to large deviations from the predicted mechanical behavior for this composite. EXPERIMENTAL PROCEDURE Epitaxial Growth of Eutectic Alloy. The A1-CuA1, eutectic alloy was prepared by an epitaxial growth process. Sections of a master alloy ingot (total impurity content <0.008 pct) were placed in an alundum boat, melted. and directionally solidified to obtain a multigrained plate 12 by 2 by 4 in. This plate was tapered at one end to mate with a seed crystal 1; in. long and 4 by $ in. square. Then the seed-plate combination was placed in an alundum boat which sat in a quartz tube passing through the center of a horizontal resistance wound tube furnace. A dried argon atmosphere was maintained. The temperature gradient in the furnace was such that the liquid-solid interface of the eutectic alloy was located near the end of the furnace and could be observed through the quartz tube. Single-crystal plates were formed by melting the material back to the midpoint of the seed crystal of the desired platelet orientation and then epitaxially growing the plate from the seed by withdrawing the alundum boat from the furnace at a constant rate. This technique was used to produce aluminum and CuA12 lamellae parallel and perpendicular to the transverse direction on the plate. Metallographic examination showed that both phases were continuous across the original liquid-solid interface. It was also possible to grow a plate from two seeds placed side by side: and, although the lamellae of one seed were oriented at 90 deg to those of the second seed, the interface between the two grains remained parallel to the growth direction along the entire ingot length. Maintenance of a straight intergranular boundary during the solidification process was possible as long as both seeds were oriented with their original growth direction parallel to the solidification direction of the plate. Eutectic plates were directionally solidified at rates of 0.2, 1.0. and 4.7 cm per hr and sectioned transversely to the solidification direction to determine the apparent inter lamellar spacing of the lamellae. Metallographic examination was also employed

Jan 1, 1970

By J. L. Carne

ORE control is a matter of planning and supervision based on a foreknowledge of the content and distribution of ore. The Inspiration orebody is predominately a copper-sulphide blanket, overlain by an oxidized zone of copper silicates, copper carbonates, and barren capping. Most of the ore is treated by leaching, and the optimum requirement of this method is the goal toward which the mining schedule is aimed. Both open-pit operations and underground mining are used. The problems of ore control at Inspiration are: 1-maintenance of metallurgical balance, 2-conservation of oxide ore, and 3-mining out of the orebody at an average grade which parallels that of the reserve figure. Prior to the fall of 1926 all of the Inspiration division ores were treated by flotation, but the mixed ores of the Live Oak-Keystone section carried too large a proportion of copper in the form of chrysocolla for satisfactory flotation. For this type of mixed ore the leaching process now in use had been worked out, and the plant was put into operation in the fall of 1926. The process is known as the ferric sulphate leach, and the leaching solvents contain both ferric sulphate and free sulphuric acid. Sulphide copper in the form of chalcocite is leached by the ferric sulphate, and the oxidized copper minerals are dissolved by the sulphuric acid. From the beginning of operations to the present time the plant has treated some 67 thousand tons of ore having an average copper content of 1.142 pet, of which 0.605 pct has been in the form of oxidized copper minerals. The plant is currently being operated at the rate of 11,500 tons per day, 7 days per week. However, this tonnage is produced in 6 days of mine operation. The average current grade is approximately 1 pct copper, of which about 50 pct is present as oxidized or acid-soluble copper. Close control of metallurgical procedures is a vital factor in operation of the plant and is in turn dependent upon careful regulation of feed. An excess of sulphide copper requires the presence of more ferric sulphate in the leaching solvents. On the other hand, an excess of ferric sulphate in the presence of a low sulphide feed seriously reduces the efficiency of electrolytic precipitation in tank house operations. This condition must be avoided at all costs. An additional factor must be considered. It is known that the sulphide content is predominant in the ore which remains in the reserve. To maintain the necessary sulphide-oxide proportion, so that the entire reserve will mine out at both a grade and a sulphide-oxide ratio that will be permissible throughout the remaining life of the property, it is obvious that the oxide reserve must be currently preserved to as great an extent as possible. Fortunately much of the oxide reserve is found in the open-pit operation. This happy circumstance adds flexibility to the procedure and makes it possible to maintain the close regulation necessary to insure best overall results, both mining and metallurgical. The two ore streams from the pit and underground operations join at the coarse-crushing plant, and proper control must be exercised before the ore reaches this point. The open pit is currently providing about 55 pet of the ore and underground operations the balance of production. Both operations are under a single superintendent so that the closely interrelated problems may be more readily controlled. The least flexible of the two producing units is the underground mine, where the mining method is block caving. Although a block of ground may be undercut and developed, characteristics of the material determine its caving action. The ore in the chutes must be drawn. The area currently mined from underground is in a zone predominately sulphide and is generally

Jan 1, 1952

By Leon Y. Sadler

Three new transfer functions relating the time response of the amount of floatable mineral in the under flow to time-dependent changes in the amount of floatable mineral in the feed to the cell are presented. Using experimental stimulus-response data obtained from the flotation of -200+325-mesh glass spheres in a rotor-stator-type cell, the new transfer functions are compared to the only transfer function to be presently found in the literature by nonlinear least-squares fitting in the frequency domain. The new mathematical model based on the amount of available bubble surface area and its rate of introduction is quite promising. The transfer function which allows for a distributed first-order flotation rate constant also gives good results. Simultaneous sodium chloride tracer and glass bead stimulus-response tests show that the mixing conditions in the cell were nearly ideal. It is felt that the new transfer functions presented offer excellent alternatives for use in the design of automatic process control equipment for the continuous mechanical froth flotation cell.

Jan 1, 1974

By Edgar Rickard

E UGAR RICKARD comes of a long line of mining men and was born at Pontgibaud, France, in 1874, where his father was then in the course of his professional work. Later his father came to California and was until his death prominently connected with milling in that state and Nevada. Edgar Rickard grew up at Berkeley, attended the University of California, graduated in 1895 from the mechanical engineering course, and remained a year longer for post-graduate work in the College of Mining. His first job was as assarer for the Mountain Copper Co., Keswick, Cal., in 1896. The following year he was assayer and surveyor for the Cornucopia Mines, Oregon. In 1898, he joined the staff of the Venture Corporation, London, and for two years made examinations for them in various parts of the world. He served as mine manager at Cripple Creek and at the Briseis tin mine, in Tasmania. From 1901 to 1905, he was general manager of the Progress Mining Co. in Lower California, but when his cousin, T. A. Rickard, bought the Mining and Scientific Press he was persuaded to come to San Francisco as business manager of that periodical. A few years later the two cousins went to London and started the Mining Magewine there as a joint venture, Edgar being the business manager. In 1915, he took over control, T. A. Rickard returning to California to devote his time to the Mining u.rbd Scientific Press. In 1917 Edgar Rickard sold out to his British associates and came home to serve in the Food Administration through the war. Rickard was one of the first of the American engineers called to his staff when Herbert Hoover began his relief work by helping stranded Americans home at the be- ginning of the war and he has been with him constantly since in the Commission for Relief in Belgium, the

Jan 1, 1929

By William Priestley

IN THE manufacture of gun forgings and other steel parts that, in service, are subject to sudden high stresses and shocks, it is most desirable to use steel possessing the greatest toughness and ductility possible without sacrifice of strength. In order to obtain this condition, it is necessary to procure steel that shows the highest possible elongation and reduction of area without lowering the tensile strength and elastic limit. Proper heat treatment of the steel can control this condition within certain limits. When heat treatment has failed to produce the desired results, metallurgists have used steels containing molybdenum, zirconium, vanadium, chromium, tungsten, etc. The purpose of this paper is to describe a method by which these desired physical properties may be procured-by the elimination of certain impurities that inherently exist in steel made by the open-hearth process, and without the use of expensive alloys. DESIGN OF GUN FORGINGS In the manufacture of gun forgings, a certain elastic limit is fixed by the designer, and the walls of the gun are made of the proper thickness, allowing a suitable factor of safety for the high stresses and sudden shocks that occur during gun firing. The elastic strength of the gun is about 1.4 times the stress set up at any point along the bore of the gun during firing with the maximum powder charge.

Jan 12, 1921

By P. W. Bakarian, C. H. Mathewson

THE greater part of the literature on the plastic behavior of magnesium dates back to that active period of research in crystal mechanics immediately following the widespread preparation of isolated metallic crystals large enough for the application of conventional methods of mechanical testing. These investigations, performed on single crystals, revealed an analogy between zinc and magnesium with respect to their fundamental mode of deformation, in the operation of glide and twinning, respectively, on crystallographically similar planes. Mathewson and Phillips (1928)1 by geometrical analysis identified the {102} ? plane in magnesium as the twinning plane, Schiebold and Siebel (1931)2 confirmed this twinning plane, established the basal plane as the plane of slip, and a digonal axis I as the slip direction. They also reported twin formation in a secondary role on the {101} plane. Schmid (1931)3 summarized and discussed the earlier evidence and presented the results of his own experiments on the effect of temperature and small amounts of soluble elements. From X-ray studies on extended wire single crystals, he concluded that slip along the basal plane in the direction of the digonal axis I closest to the stress direction was the sole slip system operating at atmospheric temperature. This conclusion was based on experiments with crystals of many different orientations relative to the stress axis. Special attention was directed to the basal plane and the digonal axis I, as the plane and direction of greatest atomic density. Determinations of resolved shear stress at the onset of plastic yielding-i.e., the critical resolved shear stress-were also made; the average of these calculations showing 82.9 grams per sq. mm. This value varied only slightly up to 300°C., at which temperature it was found to decrease to 70.5 grams per sq. mm. The extension of his crystals at temperatures above 225°C. further disclosed the presence of new slip systems. After the original specimens of circular section had been

Jan 1, 1943

By J. J. Bodmer

Although the similarity between puzzolana, or trass, and blastfurnace slag, as seen by comparison of the analyses, is a well-known fact, blast-furnace slag has not been used commercially as a substitute for those cementing materials. The reason, the writer apprehends, lies in the fact that unless such slag is disintegrated or subdivided by rolls, the process must either be too costly, or the material is not in a fit and proper condition for the purpose. In order to produce a reliable slag cement, the slag must be ground together with the lime into an impalpable powder. The subdivided slag must, therefore, be perfectly dry, and, at the same time, friable. The stronger the hydraulic properties of the lime, the more reliable the slag cement will be, and practice has proved, that the slag from a gray-iron furnace gives the best results. The slag cement which has given the results shown in the annexed table, under pressure tests, was composed of six parts of slag, from a blast-furnace producing No. 3 foundry iron, and one part of lime, of medium hydraulic properties. The above-described class of cement bears storing as long as most Portland cements, and the cheapness of its production is self-evident.

By J. J. Bodmer

ALTHOUGH the similarity between puzzolana, or trass, and blast-furnace slag, as seen by comparison of the analyses, is a well-known fact, blast-furnace slag has not been used commercially as a substitute for those cementing materials. The reason, the writer apprehends, lies in the fact that unless such slag is disintegrated or subdivided by rolls, the process must either be too costly, or the material is not in a fit and proper condition for the purpose. In order to produce a reliable slag cement, the slag must be ground together with the lime into an impalpable powder. The subdivided slag must, therefore, be perfectly dry, and, at the same time, friable. The stronger the hydraulic properties of the lime, the more reliable the slag cement will be, and practice has proved, that the slag from a gray-iron furnace gives the best results. The slag cement which has given the results shown in the annexed table, under pressure tests, was composed of six parts of slag, from a blast-furnace producing No. 3 foundry iron, and one part of lime, of medium hydraulic properties. The above-described class of cement bears storing as long as most Portland cements, and the cheapness of its production is self-evident.

Jan 1, 1874

By Charles H. Thurman

THE data given herein were first included in a paper read before San Francisco Section, A.I.M.E., in October 1940, and are applicable to conditions existing until the gold-dredging industry was temporarily suspended by Order L-208 of the War Production Board in the fall of 1942. There is no doubt that gold dredging will be resumed at some future time, and on a scale equivalent to that of the recent past, but the cost data given herein will be affected to some extent by conditions existing at that time, and by labor rates and prices for materials then prevailing. DRAGLINE VS. BUCKET-LINE Dragline dredges should not be considered for mining large areas of low-grade placer gravel. In such deposits a dragline operator with a 5-cu. yd. dragline would be doing well to maintain an operating cost of 6¢ per cu. yd., while a bucket-line dredge with buckets no smaller than 7 cu. ft. each could be operated for less than 4¢ per cu. yd. By operating costs, I mean only field charges such as direct operating and overhead, which includes pay roll, power or fuel, repairs and maintenance, insurance, land clearing and general expenses, but not including such items as prospecting, depletion, depreciation of equipment, State Franchise Tax, or Federal Income Tax. It is not to be inferred that a dragline dredge cannot make a profit in small low-grade deposits, where the extent of a given property is too small to justify the use of the more costly bucket-line dredge. Consider small deposits of 10¢ gravel. It is obvious that a deposit of 5,000,000 cu. yd. of such gravel could not be mined profitably with a new, 7-cu. ft. bucket-line dredge, costing upward of $250,000, and with operation costs at 4¢ per cu. yd., paying for land or royalties out of earnings; nor could a new combination 5-cu. yd. dragline and washing plant, worth upward of $125,000 and working at 6¢ per cu. yd., make a profit. These comparisons are made because the hourly capacity of both types of dredges is nearly the same. The bucket-line dredge averages 93 per cent operating time in 24 hr., as against about 80 per cent or less for the dragline dredge. These percentages of operating time are important factors in making a comparison of this kind. The bucket-line dredge has greater efficiency in cleaning soft bedrock, usually characteristic of dredging deposits, and a 7-cu. ft. bucket-line dredge can be efficient to a total digging depth of 80 ft. (65 ft. below water, 15 ft. above). A 5-cu. yd. dragline dredge is limited to about 30 ft. of efficient digging depth, 5 ft. of which could be above water. In small areas of gravel not over 30 ft. deep, a 5-cu. yd. dragline dredge could make a profit where a bucket-line dredge could not do so. As an example: compare the dredging of 3,000,000 cu. yd. of 10¢ gravel 28 ft. deep with bucket line, costing 4¢ per cu. yd., and with a 5-cu. yd. dragline dredge operating at 6¢ per cu. yd. Because of small yardage, new equipment is not warranted, therefore we can assume that the dredges of both types are used equipment, fully depreciated. Dismantling, mov-

Jan 1, 1945

By Charles H. Thurman

The data given herein were first included in a paper read before San Francisco Section, A.I.M.E., in October 1940, and are applicable to conditions existing until the gold-dredging industry was temporarily suspended by Order L-208 of the War Production Board in the fall of 1942. There is no doubt that gold dredging will be resumed at some future time, and on a scale equivalent to that of the recent past, but the cost data given herein will be affected to some extent by conditions existing at that time, and by labor rates and prices for materials then prevailing. Dragline vs. Bucket-line Dragline dredges should not be considered for mining large areas of low-grade placer gravel. In such deposits a dragline operator with a 5-cu. yd. dragline would be doing well to maintain an operating cost of 6¢ per cu. yd., while a bucket-line dredge with buckets no smaller than 7 cu. ft. each could be operated for less than 4¢ per cu. yd. By operating costs, I mean only field charges such as direct operating and overhead, which includes pay roll, power or fuel, repairs and maintenance, insurance, land clearing and general expenses, but not including such items as prospecting, depletion, depreciation of equipment, State Franchise Tax, or Federal Income Tax. It is not to be inferred that a dragline dredge cannot make a profit in small low-grade deposits, where the extent of a given property is too small to justify the use of the more costly bucket-line dredge. Consider small deposits of 106 gravel. It is obvious that a deposit of 5,000,000 cu. yd. of such gravel could not be mined profitably with a new 7-cu. ft, bucket-line dredge, costing upward of $250,000, and in operation costs at 4¢ per cu. yd., paying for land or royalties out of earnings; nor could a new combination 5-cu. yd. dragline and washing plant, worth upward of $125,000 and working at 6¢ Per cu. yd., make a profit. These comparisons are made because the hourly capacity of both types of dredges is nearly the same. The bucket-line dredge averages 93 per cent operating time in 24 hr., as against about 80 per cent or less for the dragline dredge. These Percentages of operating time are important factors in making a comparison of this kind- The bucket-line dredge has greater efficiency in cleaning soft bedrock, usually characteristic of dredging deposits, and a 7-cu. ft. bucket-line dredge can be efficient to a total digging depth of 80 ft- (65 ft. below water, 15 ft. above). A 5-cu. Yd. dragline dredge is limited to about 30 ft. of efficient digging depth, 5 ft. of which could be above water. In small areas of gravel not over 30 ft. deep, a 5-cu. yd. dragline dredge could make a profit where a bucket-line dredge could not do so. As an example: compare the dredging of 3,000,000 cu. yd. of 10 gravel 28 ft. deep with bucket line, costin 4¢ Per cu. yd., and with a 5-cu. yd. dragline dredge operating at 6¢ Per cu. yd. Because of small yardage, new equipment is not warranted, therefore we can assume that the dredges of both types are used equip-ment, fully depreciated. Dismantling, mov-

Jan 1, 1946

By L. Massé, Th. Koulomzine

Note on Haalck&apos;s Formula Quantitative interpretation of magnetic anomalies is admittedly a difficult process. Few authors have attempted a general approach to this problem. A number of publications can be found widely scattered through scientific periodicals, which present particular solutions obtained by the introduction of simplifying assumptions, such as the reduction to an equivalent magnet or pole, or the hypothesis of a very thin vein. The practical applications of these solutions are necessarily very restricted though they have given satisfactory results in isolated cases. Probably the most widely known equation of this type was given by Haalck.&apos; His classic work contains curves calculated for typical conditions, and has now become a standard reference. However, a close scrutiny of Haalck&apos;s derivation reveals that, aside from the hypothesis of a thin vein, it rests on the

Jan 1, 1949

By L. Massé, Th. Koulomzine

NOTE ON HAALCK&apos;S FORMULA QUANTITATIVE interpretation of magnetic anomalies is admittedly a difficult process. Few authors have attempted a general approach to this problem. A [ ] number of publications can be found widely scattered through scientific periodicals, which present particular solutions obtained by the introduction of simplifying assumptions, such as the reduction to an equivalent magnet or pole, or the hypothesis of a very thin vein. The practical applications of these solutions are necessarily very restricted though they have given satisfactory results in isolated cases. Probably the most widely known equation of this type was given by Haalck.l His classic work contains curves calculated for typical conditions, and has now become a standard reference. However, a close scrutiny of Haalck&apos;s derivation reveals that, aside from the hypothesis of a thin vein, it rests on the

Jan 1, 1947

Dr. Robert Bell died on June 18, 1917, at Rathwell, Manitoba, in the seventy-seventh year of his age. He was born in Toronto on June 3, 1841, and was a son of the Reverend Andrew Bell of the Free Church of Scotland. He was educated at the Grammar School of the County of Prescott, Ontario, and subsequently at McGill University and the University of Edinburgh. While still a boy of 15 years of age, he joined the Geological Survey of Canada, under Sir William Logan, in the capacity of Junior Assistant, and was rapidly promoted in the service, where lie remained-with but one short interval-for over fifty years, retiring as Chief Geologist and Acting Director in the year 1908. He married Agnes, daughter of the late Alexander Smith, of Westbourne, Glasgow, and Auchentroig, Stirlingshire, and is survived by one son and three daughters. Practically nothing was known of the geology of Canada, and very little with reference to its geography, when Mr. William E. (afterward Sir William) Logan was called upon to undertake the organization of the Geological Survey of Canada, and as its first Director to carry out the mapping of this vast area and the examination of its mineral resources. Dr. Bell&apos;s work as a member of Sir William Logan&apos;s staff in these early years consisted, therefore, exclusively of geological reconnaisance and geographical exploration. Even in later years, and until his administrative work confined him to his office, Dr. Bell&apos;s work remained essentially of this character. The exploratory traverses and track surveys which lie carried out were, chiefly within the area of the great Laurentian protaxis or Canadian Shield, which forms such a striking feature and which has played and will continue to play so important a role in shaping Canadian history and in influencing the character of the Canadian people. Dr. Bell&apos;s lines of exploration crossed and re-crossed this area from the Atlantic Coast to the Great Lakes on its western border, and from Lake Huron and Lake Superior on the south to the Arctic Sea, following the water ways of this great land. He also worked along the coast line when acting as geologist and naturalist to the "Neptune" Expedition in 1884, and to the "Alert" Expedition in 1885, both of which carried out explorations in Hudson&apos;s Bay and Hudson&apos;s Straits; also on the "Diana" Expedition in 1897, when he surveyed the coast of Baffin Land, and was one of the first white men who penetrated to the great &apos;lakes in the interior of that immense

Jan 11, 1917

By N. Mungan, J. L. Thompson

Laboratory displacement tests were performed to study oil recovery elficiency by gravity drainage in fractured systems under miscible conditions. The porous media used were cylindrical Berea and Blue Jacket sandstone cores containing a number of well-defined, artificially formed, vertical and subvertical fractures. Butane and Soltrol 130 were the two miscible fluids used. The purpose of the study was to examine the influences of the displacement rate, fracture density, fracture orientation, fracture permeability, matrix permeability, crossflow, core length and connate water on the oil recovery. It was found that displacement rate, matrix permeability and the subvertical fractures affected oil recovery most. The critical flow rate, based on the matrix permeability, was found to be a significant factor in the process. For displacement rates below the critical flow rate, the oil recovery efficiency appeared to be unaffected by the density of the subvertical fractures. At the high displacement rates, the fracture density becomes important, with the recovery being the most efficient in the core having the greatest number of the subvertical fractures. The magnitude of the fracture permeability, the fracture orientation, the core length and the connate water have little effect on the oil recovery efficiency. INTRODUCTION Gravity drainage under miscible conditions from relatively thick reservoirs can be a very efficient recovery process, especially at low flow rates where the gravity forces are dominant and, consequently, the adverse viscous fingers associated with the unfavorable viscosity ratio are minimized or eliminated. Such a process might involve injection of an LPG or some solvent bank at the crest and then driving the bank with dry gas. For some favorable combination of reservoir temperature, pressure and oil composition, there may be no need to inject any solvent bank, since enrichment of the dry gas by the light ends of the crude creates an in-situ solvent bank. Slobod and Howlett2 have studied the effects of gravity segregation in vertical unconsolidated porous media under miscible conditions in the laboratory. The main variables in their study were the viscosity ratio, the density differences and the rate of flow. In this study, the objective was to find the influence of fractures on gravity drainage under miscible conditions. A greater number of related variables were also studied. LABORATORY STUDY CORE DESCRIPTION Displacement tests were performed mostly in cylindrical Berea sandstone cores. In a few cases in which a much lower matrix permeability was desired, Blue Jacket sandstone cores were used. The vertical fractures were formed by cutting the cores with a thin circular blade or by parting them along a bedding plane. The vertical fracture plane always contained the axis of the cylindrical core. The subvertical fractures, cut with a saw, were inclined 45&apos; from the horizontal. The line of intersection between the planes of the vertical and subvertical fractures was parallel to the circular faces of the cores in all cases except one, as depicted by VF3HF-2 in Fig. 1. Fracture geometry of the other cores is also given in Fig. 1. Tables 1 and 2 give the physical properties of the solid and fractured cores, respectively. The suffixes and prefixes shown for each core have been used throughout the paper so that the reader may discern the fracture geometry from the core numbers.

Jan 1, 1970

By Leonard B. Gulbransen, John R. Lewis, W. Martin Fassell, J. Hugh Hamilton

T. E. Leontis (The Dow Chemical Co., Midland, Mich.)—This paper is of particular interest to me because of my own work with F. N. Rhines on the oxidation of magnesium and magnesium alloys a few years ago. The authors are to be complimented on their development of an accurate and reproducible technique for measuring ignition temperatures and on their comprehensive study of the many variables that affect the ignition temperature of magnesium. It is indeed gratifying to see that they have obtained a good correlation between ignition temperatures and the oxidation rates reported by us. The correlation is valid not only with composition within one alloy system but also between alloy systems; that is, alloying elements which effect the greatest increase in oxidation rate also produce the greatest decrease in ignition temperature. There are a few points upon which I would like to comment. In attempting to correlate ignition-temperature data, one must be sure that the same definition of this quantity is used by all investigators. It does not appear to me that such is the case in the authors&apos; comparison of their data with the theoretically calculated values of Eyring and Zwolinski. The equation derived by these investigators defines the ignition temperature, To, as the temperature at the gadoxide interface, whereas the present authors use the metal temperature as the criterion for ignition. The contradiction in the effect of oxide-scale thickness on ignition temperature between the predictions of the Eyring-Zwolinski equation and the observations reported in this paper indicate that some variable has not been taken into consideration. Could that be the geometry and size of the specimen? There is a marked difference in the type of specimen used in this investigation and that used in our work which formed the basis of Eyring and Zwol-inski&apos;s theoretical treatment. Another factor which plays an important role in ignition is the vapor pressure or the rate of vaporization. Combustion can safely be assumed to take place in the vapor phase by the reaction between vaporized magnesium and oxygen. Thus, a more accurate theoretical analysis may be made on the basis of the rate of vaporization which may be the controlling rate of the process. The effect of a large number of alloying elements on the ignition temperature has been reported in this paper, but beryllium was not included. Practical experience dictates that beryllium markedly decreases the burning tendency of magnesium. I was wondering if the authors plan to study the effect of beryllium in their future work. The authors predict that concentrations of sulphur dioxide in the furnace atmosphere greater than 5.8 pct would be expected to increase the ignition temperature to values still higher than those they measured. I would like to mention that large concentrations of sulphur dioxide markedly increase the rate of combustion of magnesium once ignition has started. Although it has been shown in the paper that the ignition temperature of magnesium in oxygen increases with increasing sulphur dioxide content up to about 1 to 2 pct, in practice relatively low-melting commercial cast alloys (AZ63A and AZ92A) are being continuously heat treated at temperatures just below the melting point in air containing 0.5 to 0.75 pct SO*. In regard to the change in color of the oxide scale observed on magnesium and magnesium alloys just prior to ignition, I would like to mention that in our work alloying elements were found to color the usually white magnesium oxide even though ignition did not occur. For example, the oxide formed on Mg-A1 alloys was gray, increasing in intensity with aluminum content in the alloy. Finally, I might suggest that the authors indicate their source of the value of 0.8 g per cc for the density of MgO as it is formed on magnesium upon oxidation at elevated temperatures. W. M. Fassell, Jr. (authors&apos; reply)—The comments by Dr. Leontis are very excellent ones and I will attempt to answer them in order. First, the problem of ignition of magnesium is a rather difficult one since many factors are involved. Concerning the comparison of the To in the Eyring-Zwolinski equation, eq 4, with the experimentally determined values, it will be noted that the calculated and experimental values of the ignition temperature in Table I are not self-consistent. In the case of the 1.78 pct A1-Mg alloy the calculated value is 49°C below the experimental value; for the 3.81 pct A1-Mg alloy, 122°C below the experimental value; for Mg with 5x10-&apos; cm film, 19°C above the experimental value; for Mg with 2x10-I cm film, 28 °C below the experimental value. Thus, if it were merely a matter of difference of location of temperature measurement the calculated ignition temperature would always be below the experimental value, the difference being due to the thermal gradient through the oxide film. The possibility of a thermal gradient in the magnesium metal must be considered. From Carslaw and Jaeger,&apos;Y t can be shown that the maximum temperature gradient that could exist between the oxide-metal interface and the center of the sample is of the order of O.Ol°C. The geometry and size of the specimen could certainly have some effect on the ignition temperature. The equation for ignition that has been proposed in reference 14 is of the following type containing terms to account for this and other factors: M dT AHv(T) =Cp--------------\-J(.T—TB) + ZAHl-M A dx where AH is the heat of reaction, v(T) is the velocity of the reaction at temperature, Cp is the heat capacity of sample, M is the mass of sample, A is the area of sample, t is time, J is the total coefficient of heat transfer outward from the reaction zone, TR is the temperature of the bath or furnace, and AH,, is the heat associated with any phase change involved. Prior to the instant of ignition, the vapor pressure of magnesium is of no special significance. After ignition, neither eq 4 nor the above equation is applicable. The actual combination of magnesium cannot safely be assumed to take place in the vapor phase. While experimental data is lacking to support a hypothesis that ignition does or does not occur in the vapor phase, some observation on the pressure ignition experiments may be of interest. At high oxygen pressures, once ignition has occurred, the reaction of magnesium with oxygen approaches near explosive violence, the entire sample being consumed in probably less than 1 sec. At atmospheric pressure it usually requires 15 to 20 sec. Thus it appears that the oxygen concentration becomes the rate determining factor. Further, if burning magnesium is observed through darkened glass (Lincoln Super-visibility Shade No. 12) the magnesium sample is very much hotter than the "smoke" and the outline of the sample is retained perfectly. No "flame" is visible above the metal. No work was done on Mg-Be alloys. We do, however, intend to study this problem in the near future.

Jan 1, 1952