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By AIME AIME

THE MILLING PROGRAM on Monday required a morning and afternoon session with a special luncheon of the Milling Committee in the Engineers Club at noon. Grinding and flotation were the main subjects of the papers which drew an attendance of fifty or more at each session. Under the direction of Chairman Arthur F. Taggart in the forenoon, all papers on flotation and general milling were considered, leaving the, after- noon free for crushing and grinding, with Dr. Galen H. Clevenger in the chair. The high spot was the first public presentation of the idea of size reduction of minerals by explosive shattering.

Jan 1, 1932

By J. C. Bowling

Factors governing the choice of building types, materials, and the details of construction for mineral processing plants in remote locations with cold climates are discussed. Alternative types of building insulation in common use are described and a comparison is made of their suitability under various conditions.

Jan 1, 1973

By Robert J. Linney

IN the latter part of the year 1943, it was decided to experiment with diamond-drill blast holes in the Old Bed magnetite mine at the Mineville mines of the Republic Steel Corporation, in sections of the ore body where vertical dimensions were large and where development for sublevel bench stopes had already been done. Other possibilities of application for this drilling method were in the reclamation of old floors and pillars in the pitching folds of the Lower Old Bed ore body and in the reclamation of large vertical pillars in the " 21 " Bonanza-Joker ore body. The cross sections and views of the two ore bodies will show the conditions to which the diamond-drill method has been applied. The Lower Old Bed magnetite ore body is lenticular and dips approximately 30° toward the southwest and strikes generally southeast and northwest. An intensely folded structure, which pitches generally on an axis approximately at right angles to the strike, is continuous from the uppermost old workings down to the present bottom of the mine, and probably continues with depth. On successively lower levels this folded structure becomes greater in vertical dimensions and the two folds are closer together in a horizontal dimension. As shown in Fig. 1, this folded structure gives to parts of the ore body considerable vertical dimensions, which provide areas that were developed for sublevel bench stoping above a slushing drift, as shown in Fig. 2. It was in one of these previously developed sublevel bench stopes, No. 26 on the 1185-ft. level, that the diamond-drill blast-hole method was applied for mining the upper sublevel. The lower sublevel was underhand-bench-drilled to the finger raises of the scram drift below. The roof of this stoping area began to break and spall on the lower sublevel, causing the roof to become inclined. Along the hanging wall of the stope this reached an elevation about half way between the two sublevels. A smooth schistose seam in the hanging wall causes the ore to break away freely. Therefore, if bench mining had been carried on in the upper sublevel, a dangerous condition of poorly supported benches would have resulted. When the drillers were slashing across the sublevel to footwall and hanging wall to make a bench, spalling roof of the opening below would probably have carried away part of the bench, and would have left a dangerous place in which to continue underhand benching. A 6 by 8-ft. raise had been driven between the lower stope and the upper sublevel about 200 ft. from a manway. It was decided to use this raise for a starting cut and to arrange holes as shown in Fig. 3, so that a slot would be opened across the ore body between the footwall and hanging wall. This slot would then be used for the purpose of breaking into the ore; by blasting vertical rings of holes drilled at right angles to the axis of the pitching folds as nearly as possible. Layouts of this slot drilling and the rings adjacent are

Jan 1, 1945

By G. E. Warner

This paper presents a review and analysis of a highly stratified Burbank sand waterflooding project in Osage County, Okla. Permeability values in this reservoir range from less than 0.1 md to nearly 3 darcys, and 90 percent of the permeability capacity is contained in approximately 26 percent of the reservoir volume. To date, the project has recovered 103 bbl/flood-acre-ft since waterflood initiation, which represents 50 percent of the total preflood cumulative and 12 percent of the original stock-tank oil in place (OSTOIP). The outstanding feature of this project has been the economical recovery of a sizable quantity of oil at water-oil ratios higher than those normally considered prohibitive. Nearly 70 percent of the recovery to date has been achieved at water-cut values greater than 95 percent. An ultimate waterflood recovery of 138 bbl/flood-acre-ft, or 67 percent of the preflood recovery and 16 percent of the OSTOIP, is predicted. The ultimate primary and secondary recovery will be 302 bbl/acre-ft and approximutely 40 percent of the OSTOIP. The actual performance is compared with the predicted performance by both the Stiles and the Dykstra-Parsons methods, and applicabiliry of these methods to this particular reservoir is evaluated. The performance of this project indicafes highly stratified reservoirs can be successfully and economically waterflooded when either reservoir or economic conditions preclude the use of mechanical and chemical means of mdbility control. Introduction The concept of the stratified reservoir has been used by reservoir engineers for many years to describe the vertical variations in permeability within a reservoir. Numerous techniques have been advanced to predict the expected waterflood performance of the stratified reservoir, and in most cases they prove reasonably accurate. The advisability of undertaking a prospective waterflooding project is determined, based on these predictions. This paper presents a review of a waterflood project in a highly stratified reservoir that normally would be considered a marginal prospect, at best, when evaluated with any prediction technique properly adjusted for reservoir conditions. This kind of project can be successful if the predictions are used to design a waterflooding system capable of economically recovering an oil cut of 1 to 2 percent. Experience gained in the early stages of the project's development was profitably applied in the later development of the entire reservoir and could prove beneficial to other projects with similar reservoir stratification or fracture characteristics. The East Burbank pool waterflooding operations have been developed in three separate stages, each of which has been a separate project: Mid-Burbank Unit, Flood A, and Stanley Waterflood. When referred to collectively, these projects will be called the Composite Project. The Skelly Oil Co.'s Barber lease — the only portion of the pool not included in the three projects — is operated as a separate facility with a cooperative injection-well agreement. It contains less than 80 productive acres and would not significantly affect the over-all performance picture.

Jan 1, 1969

By Earl E. Hunner

At the end of the year 1914, the main North Star incline shaft had reached the 6300-ft. level, and encountered a vein dipping southwest, or exactly opposite to the North Star. Subsequent development failed to find the North Star vein continuing beyond this intersection, so the 6300-ft. level proved to be the bottom level of the North Star vein, and very little ore was found below the 5300-ft. level. The new vein was called No. 1 vein. Development work carried on for the next 10 years discovered a third vein, called No. 2, branching off about 1200-ft. to the east, with a north-south strike and a west dip. No. 1 vein had about 900 ft. of backs above the 6300-ft. level before reaching the boundary line of the property, and No. 2 vein died out about 800 ft. above. During these 10 years, the money spent on development was increased from $54,000 in 1914 to $184,000 in 1924; the ore reserves steadily diminished until it was evident that the mine must close down unless by some means a large tonnage of ore could be developed quickly. Problems and Plans In order to understand what follows, the general layout and equipment of the mine as it was at this time must be briefly described. Fig. 6 indicates the relative positions of the shafts. The North Star (incline) shaft extends from the surface to the 6300-ft. level following the vein, with an average pitch of 26", between limits of 11" to 38. There is at present no hoisting equipment in this shaft between the 2700 and 4000-ft. levels. The Central (vertical) shaft connects with the North Star shaft at the 4000-ft. level, at a vertical depth of 1600 ft. In two compartments of this shaft 4-ton skips made a vertical turn of 15 ft. radius at the 4000-ft. level and ran down the incline to the 6300-ft. level. A cage operated in the third compartment to the 4000-ft. level. Men and supplies were transferred to a truck running on a third track down the incline shaft to the 6300, operated by an electric hoist on the 4000-ft. station. When the shifts changed double-deck cages were substituted for the skips in the vertical shaft, but they stopped at the 4000-ft. station. The truck to 6300 carried 22 men, and it required 8 min. to make a trip down and back. This arrangement was satisfactory as long as a fairly large proportion of the work was

Jan 1, 1930

By C. S. Finney, John Mitchell

The introduction of ovens for the production of metallurgical coke is believed to be due to L. L. Norton who operated an iron foundry in the vicinity of Connellsville, Pa. Persuaded by his foreman, an English immigrant named Nickols from Durham, L. L. Norton put up a 12-ft square oven which produced coke in 1833. The coal used was taken from a local mine at Mounts Creek. The oven seems to have been used in conjunction with the customary method of coking in mounds. It was in the Connellsville district also, in 1841, that two carpenters, Provence McCormick and James Campbell, formed a partnership with John Taylor, a stone mason, for the manufacture and sale of oven coke. The task of the mason was to construct the ovens, while the carpenters were to build the arks by which the coke could be taken by water to the market at Cincinnati. The following account of the enterprise was given by McCormick: "James Campbell and myself heard in some way that I do not now recollect that the manufacturing of coke might be made a good business. Mr. John Taylor, a stone mason, who owned the farm on which the Fayette coke works now stand, and who was mining coal in a small way, was spoken to regarding our enterprise, and proposed a partnership-he to build the ovens and make the coke and Mr. Campbell and myself to build a boat and take the coke to Cincinnati, where we heard there was a good demand. This was in 1841. Mr. Taylor built two ovens. I think they were about 10 feet in diameter. My recollection is that the charge was 80 bushels. The ovens were built in the same style as those now used, but had no iron ring at the top to prevent the brick from falling in when filling the oven with coal, nor had we any iron frames at the mouth where the coke was drawn. The top and mouth had to be repaired when they fell in. In the spring of 1842 enough coke had been made to fill two boats 90 feet long-about 800 bushels each-and we took them to Cincinnati down the Youghiogheny, Monongahela, and Ohio, but when we got there we could not sell. Mr. Campbell, who went with the boats, lay at the landing some two or three weeks, retailing out one boatload and part of the other in small lots at about 8 cents a bushel. Miles Greenwood, a foundryman of that city, offered to take the balance if he would take a small patent flour mill at $125.00 hi pay, which Mr. Campbell did. He had it shipped here. We tried it, but it was no good, and we sold it to a man in the mountains for $30.00, and thus ended our coke business." So successful did the coke subsequently prove to be in use that the three partners were asked to deliver more. Evidently they had had enough of the coke business, however, for they refused to have anything more to do with it. Few ovens were built between 1841 and 1855, and it is reported that in the latter year, "there were only 26 coke ovens along the river above Pittsburgh". Successful coke makers of these years included Mordecai Cochran, Richard Brookius, and Colonel A. M. Hill. It was the use of coke in 1859 in the Clinton furnace erected by Graff, Bennett and Co. in a plant on West Carson Street, Pittsburgh, that brought the real beginning of the coke-iron era in America. Here the successful use of Connellsville coke as a blast-furnace fuel was demonstrated beyond all possible doubt, and from the year 1859 the coking industry expanded tremendously. The era of beehive coke ovens During the latter half of the nineteenth century and the early years of the twentieth, the major percentage of metallurgical coke produced in the United States came from beehive ovens. It was not until 1893 that the first battery of by-product ovens came

Jan 1, 1961

By Milton E. Wadsworth, J. Brent Hiskey

Galvanic interaction between particulate chalcopyrite (CuFeS2) and copper results in the rapid conversion of chalcopyrite to chalcocite. The effects of temperature, surface area, concentration of sulfuric acid and agitation were systematically evaluated. The kinetics were found to be controlled by a steady-state current flow controlled by the effective anodic and cathodic surface areas involved in the galvanic couple. The experimental activation energy was 11.5 and stoichiometric data and reaction products have been characterized. The overall kinetic system has been evaluated based upon an electrochemical model.

Jan 1, 1974

By D. L. Katz

The Oklahoma City Wilcox sand, discovered on March 26, 1930, has produced 394 million barrels of crude oil and 819 billion cubic feet of natural gas as of July I, 1941. The 100,000-bbl. wells, pressures of 2600 lb. per sq. in., and flows of "Wilcox" sand caused exciting moments in the early life of the field. Today the reservoir is substantially at atmospheric pressure and produccs 75,000 bbl. of crude oil daily through 466 pumping wells. Early estimates of the initial crude-oil content of the "Wilcox" sand1 indicated that about 600 million barrels of crude oil would remain after the pumping wells ceased to produce economically. The Oklahoma City Wilcox Secondary Recovery Association was organized in March 1939, to study the possibilities of recovering a portion of this oil. Only a very brief summary of the information available can be presented in a paper of this nature. Some simplifications and generalizations are necessary to give the over-all picture in SO short a space and statements made may not apply to a specific well or small area. Since the area under consideration by the Association is that north of the south line of section 22-11-3, some information will be presented for areas A, B, C, 1, 2, 3 and part of 4 of Fig. 2, and some will be for the entire Wilcox zone west of the fault. This paper is essentially an abstract of an engineering report made by A. D. Small, B. T. Murphree, M. C. Sons and the author for the Oklahoma City Wilcox Secondary Recovery Association on Sept. 5, 1940. Some information is taken from a report to the Association by C. V. Sidwell on March I, 1940, and some data are from a recent report by A. D. Small. It gives a brief history of the Wilcox zone, estimates of the initial oil contents of the reservoir, present conditions in the reservoir, and probable results from several programs of "secondary recovery." History of Wilcox Zone The Wilcox sand lies on the flank of the general structure, as shown by cross section of Fig. I2,3,4 and areal map of Fig. 2. The sand is of Ordovician age and is composed of well assorted, rounded sand grains with no apparent cementing material. The loose texture of the sand is shown by the enormous quantities of sand production from the wells and from the soft cores. The productive sand varies from zero thickness on the east to 220 ft. maximum and to zero thickness on the west because of water. The development of the pool4 as dependent upon drilling restrictions within the city limits, which accounts for the spacing of drilling operations listed in Table I by years. In 1931, areas I through 7 of Fig. 2 were open for drilling with the exception of some west-edge properties, and all wells drilled before Feb. I, 1935, were in this portion of the field. Nearly all wells drilled from 1933 to 1037 were in areas A, B and C. Since 1939, the newly drilled wells have gradually enlarged the producing area toward the west in the northern portion.

Jan 1, 1942

By D. L. Katz

The Oklahoma City Wilcox sand, discovered on March 26, 1930, has produced 394 million barrels of crude oil and 819 billion cubic feet of natural gas as of July I, 1941. The 100,000-bbl. wells, pressures of 2600 lb. per sq. in., and flows of "Wilcox" sand caused exciting moments in the early life of the field. Today the reservoir is substantially at atmospheric pressure and produccs 75,000 bbl. of crude oil daily through 466 pumping wells. Early estimates of the initial crude-oil content of the "Wilcox" sand1 indicated that about 600 million barrels of crude oil would remain after the pumping wells ceased to produce economically. The Oklahoma City Wilcox Secondary Recovery Association was organized in March 1939, to study the possibilities of recovering a portion of this oil. Only a very brief summary of the information available can be presented in a paper of this nature. Some simplifications and generalizations are necessary to give the over-all picture in SO short a space and statements made may not apply to a specific well or small area. Since the area under consideration by the Association is that north of the south line of section 22-11-3, some information will be presented for areas A, B, C, 1, 2, 3 and part of 4 of Fig. 2, and some will be for the entire Wilcox zone west of the fault. This paper is essentially an abstract of an engineering report made by A. D. Small, B. T. Murphree, M. C. Sons and the author for the Oklahoma City Wilcox Secondary Recovery Association on Sept. 5, 1940. Some information is taken from a report to the Association by C. V. Sidwell on March I, 1940, and some data are from a recent report by A. D. Small. It gives a brief history of the Wilcox zone, estimates of the initial oil contents of the reservoir, present conditions in the reservoir, and probable results from several programs of "secondary recovery." History of Wilcox Zone The Wilcox sand lies on the flank of the general structure, as shown by cross section of Fig. I2,3,4 and areal map of Fig. 2. The sand is of Ordovician age and is composed of well assorted, rounded sand grains with no apparent cementing material. The loose texture of the sand is shown by the enormous quantities of sand production from the wells and from the soft cores. The productive sand varies from zero thickness on the east to 220 ft. maximum and to zero thickness on the west because of water. The development of the pool4 as dependent upon drilling restrictions within the city limits, which accounts for the spacing of drilling operations listed in Table I by years. In 1931, areas I through 7 of Fig. 2 were open for drilling with the exception of some west-edge properties, and all wells drilled before Feb. I, 1935, were in this portion of the field. Nearly all wells drilled from 1933 to 1037 were in areas A, B and C. Since 1939, the newly drilled wells have gradually enlarged the producing area toward the west in the northern portion.

Jan 1, 1942

By Arthur Lubinski, K. A. Blenkarn

It is explained why the bottom portion of freely suspended tubing in a pumping well buckles and straightens in succession during the pumping cycle. Field evidence of resulting rod-on-tubing wear, excessive polished rod load, and excessive horsepower are given. The possibility that buckling hastens pump wear is strongly suggested. Means for either prevention of buckling or for mini-mizing its effects are explained, their relative merits compared, and field results of their use given. These means are: tension anchors, tail pipe, sucker rod guides, and corrosion inhibitors. Charts and formulas for best use of tension anchors, tail pipe, and guides are given. Use of tension anchors, or compression anchors with a heavy tail pipe, prevents both buckling and breathing, thus improving volumetric efficiency. Without tail pipe, buckling above compression anchors precludes such improvement. INTRODUCTION It is well known that a column must be subjected to some compression in order to buckle; and also that a structural member does not buckle if subjected to a tension. When consideration is given to a tubular column, the question may be asked, "How is the buckling phenomenon affected by inside or outside pressure?' At first, it seemingly should not be affected. However, this is not so. In order to arrive at a correct deduction one should first understand the basic reasons for buckling of a column under loading*. Then he should add to the effect of that loading the effect of pressure. Following this procedure, one may discover quite unexpected phenomena. In some cases, when subjected to more pressure inside than outside, a pipe may buckle under tension. In other cases, a pipe may remain straight, although subjected to a very large compression. In 1950 and 1951 it was explained why the bottom portion of a string of pipe freely suspended in a well does not buckle even when subjected to a large compression due to buoy- ancy.2,3 It was shown that the com-monly used concept of a neutral point, as being a point in the string with neither compression nor tension, is inadequate, and a more relevant definition of the neutral point was introduced. The findings were also generalized to cases of different pressures inside and outside a pipe.3a Later, in a further generalization, allowance was made for variations of stress and pressure with depth.' The findings were applied to the new technique of hydrostatic high pressure testing of lengths of pipe in the mill.5 In 1952 it was observed that previous theoretical findings apply to tubing in pumping wells. The lower part of freely suspended tubing buckles during the upstroke portion of the pumping cycle. Later, it was ascertained that this phenomenon is responsible for wear and malfunctioning of the equipment. Various remedial measures were devised and successful field trials conducted. Only brief mentions on the subject have been published.6,7 In this paper, the phenomenon of tubing buckling is explained, field

Jan 1, 1958

By M. J. Pool, with Appendix by Larry Kaufman, R. W. Krenzer

Thermodynmic properties of the Cu-Mn system have been deterrrzined in the temperature range 973" to 1273°K by measuring the vapor pressure of manganese in equilibrium with alloys of compositions varying from 0.13 to 0.91 rnole fraction manganese using the Knudsen effusion technique. Activities of manganese calculated from the measured vapor pressures show large positive deviations. Calculated heats and entropies of mixing are positive with a maximum occurring at approximately 0. 7 mole fraction manganese. ThE investigation of thermodynamic properties of binary systems involving transition metals is of considerable interest at the present time. The incomplete subvalency electron levels of transition metals lead to unusual behavior in some alloys, and considerable information can be obtained by determining the thermodynamic characteristics of these systems and relating the thermodynamic behavior to observed physical properties. The Cu-Mn system, shown in Fig. 1,' provides an example of an alloy system in which the determination of the equilibrium phase diagram and various physical properties has received considerable attention, but very little work has been done on measurement of the thermodynamic properties. Elemental copper and manganese have comparable electronegatives, atomic sizes, and cohesive energies, but differ in crystal structure. They form a continuous series of solid solutions only at high temperatures and the structure varies in a continuous manner from fcc at the copper-rich end to fct at the manganese-rich end.' The purpose of this investigation was to measure the vapor pressure of manganese over a series of Cu-Mn alloys and from these values determine the thermodynamic properties of the binary system. Vapor pressure measurements were made in the temperature range 973" to 1273"K, remaining within the solid-solution region for each alloy studied. EXPERIMENTAL PROCEDURE The relative thermodynamic properties of solid Cu-Mn alloys were determined from vapor pressure measurements by means of the Knudsen effusion tech- nique. Several discussions of the theory can be found in the literature. In the case of copper and manganese, it is assumed that manganese is the only component subliming, so the vapor pressure of manganese over the various alloys can be determined directly by measuring the weight loss of the effusion cell. This is a valid assumption since the vapor pressure of manganese is about four orders of magnitude higher than the vapor pressure of copper in the temperature region of interest. The vapor pressure of manganese over the alloys at any given temperature is then related to the weight loss of the effusion cell per unit area per unit time. An upper limit of 1.5 x 10"s has been measured for the ratio of polyatomic to monoatomic molecules for manganese vapor in equilibrium with solid manganese at 1180°K, so it was assumed that manganese vapor is monoatomic. The effusion cells, fabricated from high-purity alumina, are 26 mm high, 18 mm OD, and have a wall thickness of 1 mm. Three different sized holes of 2.03, 2.54, and 3.18 mm diam were drilled in the lids from the inside face to minimize chipping, and the lids were then slip-cast onto the cells to provide a gas-tight seal. Alumina was selected for this particular application both because of its resistance to thermal shock and reaction with manganese vapor. In the ideal case, the orifice of the effusion cell should have a knife edge so that effusion through the orifice

Jan 1, 1970

By Pei-Yung Huang, Carle R. Hayward

IN order to understand more fully the actual state of copper lost in copper reverberatory slags, a systematic study on the various reactions between certain metallurgically important copper compounds and synthetic copper reverberatory slags was undertaken by the authors in the Massachusetts Institute of Technology from 1942 to 1945.1 This paper, being only a part of the work completed, is published first because it furnishes the first experimental evidence of the formation of copper ferrite from reactions between cuprous oxide and copper reverberatory slags. CHOICE OF SLAG COMPOSITION The choice of the composition of the synthetic reverberatory slags to be used was much deliberated. Commercial copper reverberatory slags, neglecting minor constituents, average:2.3 45 to 47 pct FeO; 38 to 42 pct Si02; 5 to 6 pct CaO; 5 to 6 pct A1203. Since the ternary system CaO-FeO-SiO2 has been much more thoroughly studied4 and is much simpler than the quaternary system CaO-FeO-Si02-A1203, it was decided to use calcium oxide only instead of calcium oxide and alumina in the synthesis of the slags. Fortunately the composition of the average copper reverberatory slag with calcium oxide substituted for alumina lies in the region of the phase diagram clarified by the investigators. A survey of the phase diagram (Fig I) suggested two compositions: (I) the com¬position near the invariant point S, which is made up of 46 pct ferrous oxide, 37 pct silica and 17 pct calcium oxide; (2) the composition near the invariant point K with 45.5 pct ferrous oxide, 43 pct silica and 11.5 pct calcium oxide. Composition S has the advantage that the percentages of ferrous oxide and of silica almost exactly coincide with those of the average reverberatory slags, and that of the lowest melting point (1093° ± 3°C) in the whole ternary system. The calcium oxide content, however, obviously is high. Since we have already substituted calcium oxide for alumina, and since calcium oxide is more basic than alumina, the resulting slag is decidedly too basic. Composition K has a slightly higher melting point (1105° ± 3°C) and a higher silica content, but this is preferred because the slag is saturated with respect to tridymite and thus should not corrode siliceous refractories. Besides, it is less basic than composition S. The final slag composition, therefore, was chosen to

Jan 1, 1947

By Robert H. Richards

Jan 1, 1897

By E. S. Leaver

THE presence of soluble base metals in precious-metal ores usually precludes cyanidation as the best method of treatment. The laboratory experiments described in this paper show the possibility of cyanidation of gold-silver ores containing less than 0.5 per cent. of cyanide-soluble copper and the effect of zinc in the cyanide solution. PART 1.-COPPER IN CYANIDATION The proposed process of cyanidation of precious-metal ores containing copper is based on the regeneration of about 80 per cent. of the cyanide used in the dissolution of the copper. The method of regeneration involves a combination of the sulfide and the acid precipitation of the copper and silver. The resultant solution is made alkaline with lime, and the gold is precipitated with zinc dust.

Jan 1, 1929

By Paul Beck

WHILE the recrystallization properties of most of the practically important metals are known in considerable detail, those of lead are still relatively little known in spite of some valuable contributions in recent years. This condition probably is due to the difficulties involved in applying the usual methods of investigation. The microscopic method of discriminating between the elongated deformed grains of the familiar lenticular shape and the equiaxed recrys-tallized grains is not successful with fully refined lead: lenticular grains cannot be seen under the microscope with any amount of deformation whatever. The probable cause of this phenomenon was recognized long ago. With large deformation the time of recrystallization at room tem-perature is shorter than the period of preparation of the specimen for microscopic examination; with small deformation the deformed grains are not noticeably elongated in the first place. However, with small deformation the migration of the grain boundaries can be observed. The only systematic investigations yet published on the recrystallization of lead, those by R. S. Russell,1-3 make use of the conditions just described. Russell considered recrystallization completed when no further migration of boundaries could be observed. A serious objection to this method arises from the fact that there is no evidence to support the view that grain-boundary migration results only from recrystallization. If such migration does result from other sources after recrystallization is com-pleted-after the deformed grains have been replaced by new unde-formed grains-the "time of recrystallization" as determined by Russell may be too long. Furthermore, the method is subject to the above-mentioned limitation because of the unavoidable period of preparation of the sample. According to Russell, even with the fastest technique 5 min. elapsed before the deformed samples were ready for examination, and consequently no recrystallization time shorter than 5 min. could be measured.

Jan 1, 1939

By M. A. Grossman

The harden ability of most steels can be predicted within 10 to I5 per cent provided the complete chemical composition is known, including "incidental" elements; and provided the as-quenched grain size is known; and provided, finally, that the composition and heating temperatures for hardening are such as to result in austenite free from carbide particles. In the method proposed herein, the steel is considered as having a base hardenability due to its carbon content alone (the hardenability of a "pure steel" of the given carbon content, without any other elements), and this base hardenability is multiplied by a multiplying factor for each chemical element present. After multiplying all these together, the final product is the hardenability. Grain size may be taken into account either in the base hardenability or after the multiplication. Hardenability is stated in terms of "ideal critical diameter", namely, the diameter of bar, in inches, that will just harden all the way through (absence of un-hardened core) in an "ideal" quench, and the calculation may also be related to the Jominy test The data bring out certain features rather clearly. For example: i. It is quite useless to attempt to predict hardenability unless all elements, including "incidentals,,) are known; thus an "incidental,, chromium content of 0.20 per cent increases the hardenability by about 50 per cent. 2. The relative effectiveness of different alloys is sometimes unexpected; thus molybdenum when calculated in this way appears to be of the same order of effectiveness as manganese, rather than much more powerful, as would appear to be common experience; observe that an increase from no manaanese to 0.20 per cent Mn provides a multiplying factor of 1.67, and an increase from no molybdenum to 0.20 per cent Mo provides a factor of 1.63, an increase of over 60 per cent in each case. However, if to a steel containing 0.50 per cent Mn there is added "20 points of manganese," the factor is raised from 2.65 (for 0.50 per cent Mn) to 3.35 (for 0.70 per cent Mn), or an increase of only 26 per cent. Thus the fist small addition of an element has a much more powerful percentage effect than an equal further addition when some is already present, and in most cases the effect of molybdenum is considered in relation to a steel in which molybdenum is absent. 3. If two elements are equally effective, a greater hardenability will be obtained by using, for example, 0.5 per cent of each than by using I.o per cent of either of them alone. 4. A knowledge of the as-quenched grain size is essential for precise work, since a difference of only one grain size number (say No. 7 instead of No. 6) makes a difference of almost 10 per cent in hardenability (in these units); this, however, does not apply to certain It should be emphasized that in chromium steels (over 0.30 per cent Cr) and chrome-molybdenum and chrome-vanadium steels, un-dissolved carbides are likely to be present in the steel as quenched, and that in such cases the charts can indicate only a maximum Passible hardenability, whereas the extent of hardening actually obtained may be much less. Thus tests on a number of chrome-molybdenum steels have indicated a degree of hardening amounting to only 50 to 65 per cent of the maximum possible, and in chromium steels from full hardening (100 per cent) down to as low as 70 per cent. On the other hand, when the amount of such elements is small (Cr under 0.30 per cent, Mo up to 0.25 per cent in the

Jan 1, 1942

By M. A. Grossman

The harden ability of most steels can be predicted within 10 to I5 per cent provided the complete chemical composition is known, including "incidental" elements; and provided the as-quenched grain size is known; and provided, finally, that the composition and heating temperatures for hardening are such as to result in austenite free from carbide particles. In the method proposed herein, the steel is considered as having a base hardenability due to its carbon content alone (the hardenability of a "pure steel" of the given carbon content, without any other elements), and this base hardenability is multiplied by a multiplying factor for each chemical element present. After multiplying all these together, the final product is the hardenability. Grain size may be taken into account either in the base hardenability or after the multiplication. Hardenability is stated in terms of "ideal critical diameter", namely, the diameter of bar, in inches, that will just harden all the way through (absence of un-hardened core) in an "ideal" quench, and the calculation may also be related to the Jominy test The data bring out certain features rather clearly. For example: i. It is quite useless to attempt to predict hardenability unless all elements, including "incidentals,,) are known; thus an "incidental,, chromium content of 0.20 per cent increases the hardenability by about 50 per cent. 2. The relative effectiveness of different alloys is sometimes unexpected; thus molybdenum when calculated in this way appears to be of the same order of effectiveness as manganese, rather than much more powerful, as would appear to be common experience; observe that an increase from no manaanese to 0.20 per cent Mn provides a multiplying factor of 1.67, and an increase from no molybdenum to 0.20 per cent Mo provides a factor of 1.63, an increase of over 60 per cent in each case. However, if to a steel containing 0.50 per cent Mn there is added "20 points of manganese," the factor is raised from 2.65 (for 0.50 per cent Mn) to 3.35 (for 0.70 per cent Mn), or an increase of only 26 per cent. Thus the fist small addition of an element has a much more powerful percentage effect than an equal further addition when some is already present, and in most cases the effect of molybdenum is considered in relation to a steel in which molybdenum is absent. 3. If two elements are equally effective, a greater hardenability will be obtained by using, for example, 0.5 per cent of each than by using I.o per cent of either of them alone. 4. A knowledge of the as-quenched grain size is essential for precise work, since a difference of only one grain size number (say No. 7 instead of No. 6) makes a difference of almost 10 per cent in hardenability (in these units); this, however, does not apply to certain It should be emphasized that in chromium steels (over 0.30 per cent Cr) and chrome-molybdenum and chrome-vanadium steels, un-dissolved carbides are likely to be present in the steel as quenched, and that in such cases the charts can indicate only a maximum Passible hardenability, whereas the extent of hardening actually obtained may be much less. Thus tests on a number of chrome-molybdenum steels have indicated a degree of hardening amounting to only 50 to 65 per cent of the maximum possible, and in chromium steels from full hardening (100 per cent) down to as low as 70 per cent. On the other hand, when the amount of such elements is small (Cr under 0.30 per cent, Mo up to 0.25 per cent in the

Jan 1, 1942

By James T. B. Ives

The map exhibited* as an example of my method of construct ing geological strata-maps is essentially an educational appliance. The method, however, is available for the production of maps of comparatively small areas, which, I am confident, will be found of great practical value to mining engineers. Before stating how the method may be so applied I will briefly describe the map exhibited, as the easiest way of making my meaning clear. Almost any region might be taken. This map represents the eastern half of the United States; and the geology is that of the map issued by the Geological Survey, and revised and extended subsequently by Professor Hitchcock for your Institute. I use ten cards, of the colors used in the geological map, each having the geog raphy of the area printed upon it; and I superimpose these in the order in which the colors occur in the index to the map. I next cut away a portion of each card, excepting the lowest, leaving only so much of each as represents the exposure of the formation of that color plus all those which overlie it. When that is done, the whole, looked at from a distance, appears to be an ordinary geological map. Looked at more closely, however, the formations are seen to overlie one another, as in nature. The cards are covered with glass in a frame, which, hinged at the bottom edge, can be let down. The cards are then found to be all hinged in the same manner, and they, too, may be lowered, one after the-other. I am told by teachers of geology of great experience and the highest reputation that by this appliance they will be assisted in im parting to students the first facts and principles of geological sci ence. Having let down the whole series of cards, until that repre senting the Archaean is seen underlying all the rest, the teacher will gradually rebuild it. The Cambrian will thus be laid down upon the Archaean before the eyes of the class, the Silurian upon that,

Jan 1, 1888

By E. C. Harder

BAUXITE is known mainly as the ore from which aluminum is smelted but it has large use also in the manufacture of artificial abrasives and as a basis for certain chemical industries. A small amount is used for refractories and for other purposes. COMPOSITION Dana and others give the mineral formula of bauxite as A1203.2H20 and the composition as A12037 73.9 pct; H2O, 26.1 pct. It has been definitely shown, however, that bauxite does not exist as a specific mineral but rather as a rock. The term is now used synonymously with aluminum ore and it embraces gibbsite (hydrargillite or alpha trihydrate), A1203.3H20 (A12037 65.4 pct; H2O, 34.6 pct); boehmite (alpha mono¬hydrate), A1203.H20(Al203) 85 pct; H2O, 15 pct); diaspore (beta monohydrate), A1203.H2O(A12037 85 pct; H2O, 15 pct), and mixtures in various proportions of any two of them. Bauxite of the Mesozoic and Cenozoic in Europe is predominantly a mixture of gibbsite and boehmite, subordinately a mixture of boehmite and diaspore, or gibbsite and diaspore. The bauxite of North and South America, tropical Africa, Asia, and Australasia consists largely of gibbsite. Corundum (A1203) is not included under bauxite, although gradations from bauxite to emery or corundum exist. All bauxite, irrespective of the aluminum minerals composing it, contains certain impurities, including silica, in the form of clay minerals (kaolinite, halloysite, and others) or quartz, iron oxide (as hematite or goethite), titania (as leucoxene or rutile), iron sulphide (as pyrite or marcasite), iron carbonate (siderite), and calcium carbonate (as calcite); the last three mentioned in minor amount and of local occurrence only. Bauxite belongs to a group of partly consolidated materials called laterites, formed by surface weathering. This group includes also impure siliceous and ferruginous bauxite and siliceous and

Jan 1, 1949

By Charles E. Glass

A steadily growing body of evidence indicates that earthquake ground motions can cause failure of rock slopes that are otherwise stable under static loading conditions. As a result, the economic optimum working slope angle is reduced when earthquake shaking is considered in surface mine design, especially in regions of high seismicity. When designing surface mine slopes to withstand seismic motion it is as important to consider the probability of experiencing various levels of ground shaking as it is to consider the characteristics of the ground motion. For most slope designs, simple probabilistic techniques such as extreme value or maximum likelihood techniques are appropriate. These techniques, however, should be applied in such a way that a probability value is assigned to characteristic ground motion parameters at the mine rather than to nearby earthquake magnitudes. Several techniques have been developed to analyze rock slopes subjected to dynamic loads. Most of these techniques do not adequately account for the ground motion parameters of concern. The best technique appears to be one that utilizes actual earthquake time histories. A simple technique that calculates block displacement for any input accelerogram, using a linear acceleration approximation between time steps, has been developed that accurately accounts for rock block response. Results using this technique indicate that the following factors are important for assessing rock slope stability in earthquake regions:

Jan 1, 1983