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By Charles T. Malcomson

In the United States, improvements in methods of combustion have made possible the use of the smaller sizes of anthracite. This coal is now being reclaimed from the culm banks accumulated by the miners in the more prodigal years. To-day the freshly mined "slush" is segrega,ted and stored against the time when, through briquetting or some other means, it may be utilized as a commercial fuel. For many years Europe has achieved enviable results in the conservation of her small coal by briquetting. This profitable industry, first placed on a commercial basis in France in 1842, has grown to such magnitude that the production during 1913 exceeded 36,000,000 short tons. Briquets produced in Great Britain are used principally for steam purposes and the greater part of the output is either exported or taken by the Admiralty. On the Continent the briquets are sold in all branches of the coal trade. The French railways consume from 16 to 40 per cent. of their fuel in briquet form, and in Germany the lignite briquets (of which 20,000,000 tons were produced in 1913) are as familiar and interesting to the American traveler as the huge porcelain stoves in which they are burned. It is not strange, therefore, that aggressive and far-sighted promoters heralded this means of converting our ubiquitous coal waste into revenue. To the superficial student the briquetting process seemed so simple as to make success easy. It is for this reason principally that the path of the industry in America has been strewn with failures. The rash enthusiast had either not worked out the mechanical problems involved or had failed to grasp the local conditions under which the briquets must be made and sold. It may be confidently stated, however, that in spite of the many vicissitudes through which it has passed, coal briquetting is now an accomplished economic success in this country, and as a profitable adjunct to the coal industry it has come to stay. Several valuable contributions relating to coal briquetting are found in the Transactions. Charles Dorrance, Jr., has sketched briefly the history of the anthracite briquetting plants from the time of Loiseau's pioneer plant down to that of the Lehigh Coal & Navigation Co. at Lans-ford. Since then this phase of the industry has remained practically at

Jan 1, 1915

By C. S. Finney, John Mitchell

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

Jan 1, 1961

By F. B. Foley

IT is with an unusual degree of personal satisfaction that I find myself in a position to pay tribute to the memory of Henry Marion Howe. One could not have spent any length of time in the presence of Dr. Howe without profiting intellectually. I do not hesitate to say that my years of association with him were the most stimulating of all my years of metallurgical study. He always impressed one with his own eager search for knowledge. Never dependent on the written word alone he sought information from anyone whose opinion he valued. His ability to piece together bits of apparently conflicting data from various sources so as to build up a logical hypothesis was unique. A small slight man with the intellect of genius he would have gone to the top in any endeavor he undertook and it was fortunate for metallurgy that its study attracted him as it did. Howe was a modest man. I like to recall an incident which left a lasting impression with me. The lab- oratory in his home, "Green Peace," was in the basement and access was by way of stairs leading down from his secretary's room. In using these stairs my attention was attracted many times to an ordinary cardboard shoe box which reposed on the shelf above the landing. It was crudely labeled on its side in black ink "Vanity Box." My curiosity was aroused to the point that I finally asked his secretary what it was. With some amusement she took the box from its shelf on the stairway, opened it and showed me its contents—numerous letters, from the foremost men of science of every civilized country throughout the world, commending his "Metallography of Steel and Cast Iron." I recall that in reply to one who thought there was not enough of Howe's own researches in his book he wrote, "Primarily I am a writer, secondarily an investigator." Howe wrote to make his readers think. No one ever strove harder than he to be right but above all, whether his viewpoint proved ultimately to be right or wrong, he was always content if by his stand, he provoked a reader to take the next step along the path to greater knowledge. I doubt that he was ever afraid to be wrong for he was always secure in the thought that his effort was guided by a sincere search for the truth. One continually searching for truth is entitled to occasional excursions up the wrong alley. A glance backward to the metallurgical confusion of some thirty to forty years ago, or need one go back so far, provides convincing proof of what a host of companions one may have in a common acceptance of ideas which the future will prove to be wholly untenable. Well over a hundred years have passed since investigators have interested themselves in the effect of increasing the temperature of iron and steel on their mechanical properties. We are told by Charles Walrandl in "Industrial Annals" for June 11, 1822, that bend tests, conducted in a Russian steel works of Prince Demidoff, on steel bars "highly heated" and bent during cooling became brittle when bent at an iris blue color. He concluded, "That when steel is heated to a temperature between 473°F and 662°F the mettle was more brittle between these limits than at a much lower or at a much higher temperature." It was a curious bit of information recognized as true to this day and still not explained satisfactorily. In 1878 Charles Houston in Annales de Mines associated this brittleness with an increase in tensile strength at 572°F. When this relatively low temperature is exceeded it is recognized that steel becomes weaker as temperature is increased up to the melting point, where no strength of practical importance remains. It is easy to believe, in fact it goes without saying, that this weakening as the temperature of a metal is increased is the result of the motion of the atoms making up the metal, a motion which itself is evidence of the temperature increase. However, if we are considering iron or steel we find that this decrease in strength is not a steady one, for, besides the increase in strength just referred to, one comes to a temperature, the critical temperature, where results of tests indicate the metal to be extremely weak and then as temperature increases to become sensibly stronger again. This apparent anomaly was made the subject of the first Howe Memorial Lecture, delivered by Albert Sauveur in 1924. It may be of interest to review these findings of Sauveur. He used two methods of investigation. One involved the twisting of bars. The bars were heated in an

Jan 1, 1951

By D. Y. F. Lai, R. J. Borg

Tracer diffusion of Fe59 has been measured in the a-stabilized Fe-1.8 at. pet V alloy from 700° to 1500°C. The activation energies are obtained in both the presence and absence of magnetic order. Furthermore, it is established that diffusion in the alloy is identical to that in pure iron and consequently the values of Do and Q accurately represent the temperature dependence for self-diffusion. The purpose of this investigation is to obtain an accurate estimate of the temperature dependence for self-diffusion in bee iron in both the presence and absence of magnetic order, and, in so doing, to establish the temperature range of the magnetic effect.&apos;" As the temperature interval suitable for diffusion measurements is severely limited in both bee phases of pure iron because of the intervening fee ? phase, the experiments were performed on an a-stabilized alloy containing 1.8 at. pet V. This alloy is bee over the entire range from room temperature to the melting point. Although there have been several independent investigations of self-diffusion of iron in a, iron,1, 3-6 there still exists considerable disagreement regarding the values of Do and Q for the paramagnetic region. The two systematic studies of diffusion in 6 iron6, 7 previously reported are also only in fair agreement; but in view of the extremely small temperature range available for diffusion studies, i.e., 1390o to 1535oC, this is not surprising. It is comparatively easy to obtain accurate values of Do and Q for the a-stabilized alloy inasmuch as measurements can be made over the entire temperature range -700o to 1500°C. However, in order to assume that these same values apply to pure iron requires careful comparison of the data in the a, and 6 regions in both the alloy and pure iron. We have made several measurements in the appropriate temperature ranges and are unable to establish any systematic difference between the diffusion coefficients of iron in pure iron and in the alloy. We therefore conclude that the values obtained for the alloy are truly applicable to pure iron; the complete evidence favoring this conclusion will be discussed later in this paper. EXPERIMENTAL The experimental methods will be given here only in barest detail since they have been thoroughly de- scribed elsewhere.l, 7 The alloy was prepared by induction melting and chill casting under argon. Diffusion samples were machined from the ingot and annealed in hydrogen for several days at 900°C to give an average grain diameter of 1 to 2 mm. The penetration profiles of the tracer were established by a sectioning technique, the residual activity being counted after the removal of each section. The tracer used is Fe59 which emits two high-energy ? rays of 1.098 and 1.289 Mev, respectively; these were detected by a ? scintillation counter equipped with a pulse-height analyzer. For the measurements in the temperature range -700o to 1130°C the samples were vapor-plated with Fe59, encapsulated in quartz under vacuo, and annealed in resistance-heated furnaces which are controlled to ±1°C. The specimens diffused at higher temperatures are prepared as edge-welded couples, the two halves being separated by a thin washer of the alloy to prevent sintering. The diffusion anneal is then carried out by inductive heating under a dynamic vacuum. The temperature is monitored pyrometrically. RESULTS The diffusion coefficients are obtained from the penetration profiles in the usual way using the error-function complement relationship. The results over the entire temperature range are shown in Fig. 1. In the linear region, 900o < t > 1500°C the least mean squares (lms) values of the diffusion coefficients are given by D = 1.39 exp[-(56.5 ± 1) x 103/Rt] cm2/sec [l] The average departure of the measured diffusion coefficients from the values given by Eq. [1]In order to determine whether or not the slope is truly constant over the entire range from 900" to 1500°C, the data are arbitrarily divided into two groups, the first containing values between -900" and 1133°C and the second between -1133o and 1500°C. The lms values for the two groups are given by Eqs. [2] and [3]: D = 0.519 exp[-55.7 x 103 /RT](900° to 1133°C) cm2/sec [2] D = 1.45 exp[-56.7x103/RT] (1133° to 1500=C) cm2/sec [3] Thus, there is no significant difference between the high- and low-temperature segments of the linear region. This not only assures us of the consistency of the values obtained by induction heating as compared to those obtained from the resistance-heated

Jan 1, 1965

By H. Ewaschuk

"INTRODUCTIONRio Algom's Stanleigh Mill in Elliot Lake began production in June of 1983. The highly instrumented mill treats uranium ore at an average rate of 3900 metric tons per day. To help ensure a good start-up and low operating costs, an extensive training program was undertaken prior to and during start-up. This presentation describes how an unconventional operating crew were combined to form an operating unit with a high efficiency in terms of production and labour utilization. I would like to point out at the outset that the operating philosophy, and the role of the supervisors and the operators as set up at Stanleigh is not unique. Indeed, I have seen the same results at a couple of small plants where efficiency meant survival. However, these situations appeared to be the exception, and it was not readily apparent how similar results could be obtained in a larger plant. A factor that existed at the time was low labour turnover. As in most parts of Canada, jobs were easy to get in the Elliot Lake camp in the late 1970's. This, and other factors, contributed to an unusually high labour turnover. This changed dramatically in the early 1980's. Indeed, Rio's two other operating mills, Quirke and Panel reduced crew sizes at this time to reduce operating costs. As a result all of Stanleigh's operators came from Quirke and Panel with experience varying from 4 to 7 years. The experience that operators had was obviously of benefit and served to reduce the scope of the training program from what would have been required with inexperienced personnel."

Jan 1, 1985

By H. S. George

A method of illumination that enhances the value of microscopic study of opaque materials, as in metallography. By simulating natural lighting, structures. That possess relief are given a natural appearance. An example is included of the revelation of a hitherto invisible microconstituent. Erroneous interpretation is avoided. For example, etching figures in ferrite, heretofore considered to be pits, are revealed as pyramids, standing in bas-relief. Suggestions on technique are given, for, while the method is obvious, practice and attention to detail are requisite. ON one occasion when it seemed desirable to reveal under the microscope not only the metallic structure of an alloy, but certain small non-metallic inclusions, that are ordinarily obliterated by etching, it was discovered that, by partly eclipsing the illuminating pencil of light rays, the residual light was thrown on the object from one side and the structure was brought out in relief. At the same time, the non-metallic inclusions were retained in clear view. In this first demonstration of lighting obliquely through the objective a card was thrust part way into the light beam between the arc iris, and the light filter; the shadow of the card on the filter served to locate its position. Next several stiff-paper disks of various diameters were mounted on flexible wires some plasticine on the other end of the wires made it easy to adjust and maintain a disk at any desired position in the light beam. To facilitate manipulation, an assortment of disks, was mounted on a wheel, as shown in Fig. 1. The purpose of partly eclipsing the light beam was discussed in an earlier paper.1 There it was pointed out that the usual methods of illuminating opaque objects for microscopic examination lend an, un-natural appearance to, the view, while the new method, which has been termed conical illumination, revels the true relationship of constituents by simulating natural lighting.

Jan 3, 1924

By A. M. Turkalo

IT is a well-known fact that martensitic steels show a greater resistance to brittle fracture than do pearlitic and bainitic steels. It was, therefore, thought worthwhile to investigate the mode of brittle fracture with respect to structure by studying the effect of various austenite decomposition products on the propagation of brittle fracture in steel by means of electron microscopy. Newly developed replication techniques together with the advantages of the electron microscope such as great depth of field, available high magnification and easy adaptation to stereophotography make electron microscopy very suitable for fracture studies. EXPERIMENTAL PROCEDURE An essentially plain carbon steel containing 0.56 pct C, 1.30 pct Mn, 0.02 pct P, 0.03 pct S and 0.22 pct Si was used in this investigation. The following table gives the heat treatments used to obtain the various austenite decomposition products: The resulting austenite grain size from the 870°C austenitization treatment was about an ASTM No. 7 grain size. The specimens which were 2 in. long, 1/4-in. diam bars were austenitized in a tube furnace through which argon gas was blown. The tube was not, however, air-free and the specimens were not completely oxide-free. Isothermal treatments were done in a lead bath. With the exception of the fully hardened specimen and the one as-isothermally transformed at 300°C the heat treated specimens were notched in the following fashion with a jeweler&apos;s saw. A cut about 0.08 in. deep was made perpendicular to the length of the bar. The specimen then was rotated about 45 deg and another cut about 0.08 in. deep was made. In the case of the hard martensitic and bainitic specimens, only one cut was made and the depth was not controlled. The design of the two-cut notch caused fracture to initiate at the point of intersection of the two saw cuts. The specimens were broken at -196°C in a Charpy impact testing machine. In order to hold the specimens securely in the impact machine two square pieces of mild steel were made about 5/8 in. long having the same cross section as that of a standard Charpy specimen but containing a hole 1/4 in. diam through the center of the cross-section. These adapters were slipped over the ends of the test specimen which was then secured tightly by means of set screws in the adapters. The specimen with the adapters was cooled to -196°C in liquid N,, then placed in the impact machine and broken within 3 sec. One half of the broken specimen was used for the electron microscope study of the actual fracture surface. The other half was nickel-plated and a cross section through the notch and fracture was polished for electron microscope examination. Carbon replicas of the fracture specimens for the electron microscope study were made essentially according to Bradley&apos;s evaporation technique.&apos; Both direct carbon and two-stage carbon replication techniques were used. In the case of the direct carbon technique, a thin layer of carbon was evaporated first at an angle of about 45 deg to the mean fracture surface and then another film of carbon was evaporated normally to the fracture surface. The carbon replica was freed electrolytically.2 The specimen was made an anode in a 10 pct Nital polishing solution. It was etched intermittently by shutting off the power now and then. The freed replica was washed in 40 to 50 pct water solution of nitric acid for 10 min or so, then washed in water and picked up on a screen for examination. The two-stage carbon replication technique involved first making a primary replica of cellulose * acetate from which a carbon replica was then made. One side of a strip of cellulose acetate wet with acetone was pressed lightly against the surface of the specimen, allowed to dry and then stripped. Prior to the deposition of carbon, which was done at 90 deg to the cellulose acetate replica surface, chromium was evaporated at 45 deg. After the carbon evaporation the cellulose acetate was dissolved in acetone according to the Jaffe method3 leaving the carbon replica preshadowed with chromium for examination.

Jan 1, 1961

By R. E. Zimmerman, Michael Sokaski, E. R. Palowitch, M. R. Geer, H. F. Yancey, W. Deurbrouck, S. C. Sun

PART 1: DENSE MEDIUM SEPARATION by M. SOKASKI, M. R. GEER and H. F. YANCEY INTRODUCTION In the early days of coarse-coal treatment by the dense-medium process in Europe, loess was one of the materials used for medium solids. Newly developed cyclones were used to reclaim and thicken :he loess suspension. On occasion, when the cyclone thickener at the Maurits mine in the Netherlands plugged, the overflow was found to be filled with clean coal free of impurity. From this observation stemmed the concept of the cyclone as a cleaning device. Development took place during World War 11 in the laboratory of the Dutch State Mines under the direction of M. G. Driessen. A 15-ton-per-hour pilot plant was in operation in the laboratory by 1945, when, in September, Driessen made the first public disclosure of the process in a paper presented to the Institute of Fuel in London. In April 1945,2 operation of this pilot plant was observed by one of the authors, who was so impressed with the potential of the cyclone process that, on his return to the United States, a research program on the cyclone was started by the Bureau of Mines immediately. This work led to an AIME paper in 1946 in which the first details of cyclone operation on American coals was made available. Commercial adoption of the cyclone proceeded in Europe, initially in the Netherlands, Germany, and France. In the United States, where coal is less costly and most of the coals are so easy to clean they respond well to simpler cleaning systems, the first cyclone plant was not built until 1961.3 As of the spring of 1967, a total of 31 commercial cy- clone plants had been installed or were under construction in the United States; this does not include eight Dynawhirlpool plants. The capacity of the cyclone plants ranges from 50 to 1300 tons per hour and totals 6600. THEORY Before discussing the theory of a dense medium cyclone, a brief description may be helpful to some readers who are not familiar with its basic operation. In a typical dense-medium cyclone, illustrated in figure 10-1, the mixture of medium and raw coal enters tangentially near the top of the cylindrical section, thus forming a strong vertical flow. The refuse moves along the wall of the cyclone and is discharged through the under- flow orifice. The washed coal moves toward the longitudinal axis of the cyclone and passes through the vortex finder to the overflow chamber.

Jan 1, 1968

The initial results from the magnetic sludge factorial experiment are summarized in Table 2. For ease in interpretation, the four separate variables are identified with the high and low level of each variable being represented by + and - signs, respectively. Several important relationships are evident from an inspection of the data in Table 2. For example, magnetic sludges were obtained only during the high temperature tests. Secondly, with the exception of the last three tests, there is an apparent correlation between settled sludge volume (or sludge settling rate) and pH. Data on residual dissolved Fee+ indicate that precipitation of iron is not complete at pH 8 and room temperature (Tests l, 3, 5, and 7). Although this should not have any confounding effect on the interpretation of the results, it is an important observation and an aspect of the overall conceptual process that should be considered carefully during any future developmental work. The data in Table 2 were employed more fully in the calculation of main effects and variable interactions for each of the three responses. For this purpose, a codified design and calculation matrix, shown in Table 3, was used. The derivation of this matrix is described by Davies.(9) The positive or negative signs for variable interactions are obtained simply by algebraic multiplication of the signs for each single variable involved in the interaction. Main effects and interactions are calculated by multiplying the response values (Table 2) by the corresponding effect elements (Table 3), summing the results, and dividing by 8. The resulting value for single variables is also sometimes designated as an average effect, since (for tyro levels of the variable only) it is simply the difference between the average response of all trials carried out at the first level of the variable (factor) and that of all trials at the second level. If the effect of one variable changes at different levels of another variable, the two variables are said to interact. The relative importance of the variable or variables in question is reflected by the absolute magnitude of the calculated main effect or interaction. First-order (two variable) interactions are usually easy to interpret in terms of the physical behavior of the system. Second- and higher-order interactions are often less readily understood, and are often of minor significance in any case. The main effects and interactions derived from the present study in the manner described above are listed in Table 4. It should be noted that

Jan 1, 1971

By H. A. Bourne, E. L. Haden, D. R. Reichmuth

A circumferential-toothed bit with novel tooth form gave improved penetration performance. In this design the exterior flank of all teeth were vertical when in rolling contact with the hole bottom. Rock chips were generated by the interior flank of the tooth displacing the rock inwardly and downslope toward the center of the hole. A unique two-cone laboratory bit assembly enabled evaluation of numerous cone and tooth configurations. Some of the variables investigated, in addition to weight on bit, rotary speed and rock type, were tooth interference, percent tooth, hole bottom angle, attack angle and relief angle. Most tests were conducted dry on a brittle synthetic sandsone or a ductile quarried limestone. Tooth configurations were found to be more significant in the ductile material. This was attributed to the deeper tooth penetration before rock failure. These studies showed that the attack angle (angle beween interior flank of the tooth and rock surface) was the controlling variable; changing the tooth configuration from the assymetric or semi-wedge to the more conventional symmetric or wedge form reduced penetration performance; and penetration performance of circumferential-type cutters was directly proportional to rotary speeds up to 200 rpm. INTRODUCTION Much of the published literature on rock-chisel interactions describe experiments wherein symmetrical wedges are vertically loaded or impacted against a smooth rock surface.1-6 are is usually taken to insure that the indentation is not made near the edge of the rock specimen less erroneous data be obtained. The literature describes relatively few studies in which the investigator deliberately attempted to take advantage of an edge or free surface. In contrast, anyone who chips ice or breaks up a concrete sidewalk almost always works near an edge. Chisel "indexing," which has been considered by some investigator1,2,6,7 makes limited application of an edge or free surface. Probably the best documented investigation into applying this idea to drilling was that of Drilling Research Inc. at Battelle Memorial Institute.&apos; Their "annular wing" percussion bit consisted of paired asymmetric chisels oriented so as to produce and move chips to the center of the hole. They predicted that the lowest energy requirement for chip generation would be achieved with a stepped hole bottom having a median angle of 45" to the horizontal. Results from limited tests showed that approximately 50 percent of the rock fragments were large and semi-circular in shape, as would be expected by a chisel impact near an edge. The remaining 50 percent were fine chips produced by the chisels in re-establishing the steps or ledges. Initial penetration rates with this bit were high, but they rapidly decreased. This was the result of excessive tooth wear caused by the constant friction on the gauge surfaces. The basic idea — circumferentially placed asymmetric chisels — still appears to have merit. If the concept could be applied to a rolling cutter bit, two of the shortcomings of the fixed chisel design could be overcome: (1) reduction in tooth friction, and (2) greatly increased cutter surface. Adapting asymmetric chisels to cutters rolling on an inclined hole bottom is restricted by bit geometry. The basic elements of roller rock-bit construction prevents the practical attainment of a 45" hole bottom angle. Nonetheless, experimentally it was considered desirable to investigate the influence of hole bottom angle to at least 40". This paper describes the laboratory studies conducted in evaluating the circumferential-toothed roller cutter rock bit. EXPERIMENTAL APPARATUS AND PROCEDURE BIT ASSEMBLY The cost of constructing a sufficient number of conventional three-cone rock bits to investigate circumferential cutter performance was prohibitive. Instead, a novel two-cone laboratory assembly which used an external bearing system was designed and constructed. The external bearings made it possible to alter the journal bearing angles and thus allow a wide flexibility in cutter configuration. Fig. 1 shows the laboratory bit assembly, the various bearing mount plates and the appropriate roller cutters for drilling shallow holes having hole bottom angles of 0, 10, 20, 30 or 40". The bit was limited to a drilling depth of 1 1/2 in. at the gauge teeth and a hole diameter of 43/4 in. This more or less intermediate size bit was chosen because it gave a more realistic match between bit teeth and the rock than would a microbit. Also, the rock sample size required was convenient and easy to obtain. CIRCUMFERENTIAL CUTTERS The tooth configuration used in our initial studies is shown in the upper half of Fig. 2. All cutters used in this series had the same tooth form — 43" included tooth angle, 2" positive relief angle and a horizontal tooth flat width of 1/32 in. Each cone cuts alternate rows except for the gauge row. The row-to-row spacing in view was 1/4 in. Static loading tests conducted earlier with asymmetrical chisels had been used to establish this spacing. These tests showed energy requirements for chip production increasing rapidly as the distances to the edge increased beyond

By H. M. Fisher, R. E. Snow, S. C. Sun

SEVERAL constructive and fundamental studies have been made in the analysis of data obtained from experiments carried on with batch ball and rod mills. The operating characteristics of ball milling in small continuous circuits have also been appraised. It is from these analyses that some of the theories of comminution have been developed. Relatively few studies of continuous rod milling have added significantly to the fundamental concepts, because seldom have they yielded sufficiently consistent results. Perhaps they have been too limited in their scope. Careful control of the variables in batch grinding is simple compared with that encountered in a continuous operation. This factor alone has discouraged many investigators. Occasionally results of systematic changes made in industrial rod mill circuits have been published, but usually the data are sketchy and are restricted because of the unwieldiness of the equipment used. The work, in general, has not been comprehensive; nevertheless it has provided empirical relationships that have bridged the gap between postulate and practice so that by proper manipulation of formulae, a mill designer can anticipate mill size and power requirements.14 Although operating variables of a small continuous mill are not so easy to control as with the batch mill, with present day devices, and with careful experimental work, consistent results can be obtained. Nearly four years ago, in the Process Laboratory, Allis-Chalmers Mfg. Co. began a systematic study of the effects of several variables upon the performance of the pilot rod mill. A mill was built in the laboratory to provide the versatility required for the proposed study. It was constructed in sections so that it could be operated, with a few modifications, as a rod mill 30 in. x 8 ft or 30 in. x 4 ft. The discharge end of the shell was flanged so that either an end peripheral discharge or an overflow discharge could be installed. Thus the performance of at least four types of mills could be studied merely by changing the type of discharge or the length of the mill shell. The grinding experiments were designed so that a study could be made of the way in which the mill speed, feed rate, and pulp density influenced the performance of both overflow and end peripheral discharge rod mills. Four sets of experimental data were collected from the four mill arrangements. The mill in each set of experiments was fed at four rates of feed depending on the length of the mill, at four pulp densities, and at five percentages of critical speed. Electrical and mechanical controls were in- stalled to regulate all these independent variables, and auxiliary devices were used to verify the precision of the controls at each point. The dependent variables used to quantify the experiments were the reduction ratio and the hew surface area produced as calculated from sieve analyses. These were incorporated with the energy factor by the calculation of both the new surface produced per unit of energy and the Bond work index.&apos; Rod wear, as a dependent variable, was not studied because of the short period of operation for each run. Exclusive of repeat runs, each set of experiments yielded 80 products, and the total study at least 320 products, all of which were quantified. With the operating information collected, these data presented a bewildering accumulation. Statistical analysis has been invaluable in unraveling the confusion and in presenting a means of establishing the nature and the magnitude of the significant variables. Data presented in this paper are those from the 30 in. x 4 ft end peripheral discharge rod mill, Fig. 1, when limestone was ground at feed rates of 1000, 2000, 4000, and 5000 lb per hr, at pulp densities of 50, 60, 70, and 80 pct solids, and at mill speeds of 50, 60, 70, 80, and 90 pct of the critical speed. These 80 tests have all been run at least twice, and occasionally a third time, to prove that the data obtained were reproducible. The techniques of operation and the methods of quantification of results are described in the following pages and the results analyzed statistically to show the significant variables. The variables are plotted to show the relationships that exist. A massive dolomitic limestone from Waukesha Lime and Stone Co. was used for feed during these experiments because of its availability and its tex-tural uniformity. This limestone analyzed 28.7 pct CaO, 21.0 pct MgO, 6.0 pct SiO2, 0.4 pct A1²O³, and 0.3 pct Fe²O³ and had a loss on ignition of 44.1 pct. It had a rod mill grindability at 14 mesh of 9.6 grams per revolution from which a work index of 13.9 was calculated. The ball mill grindability at

Jan 1, 1955

By J. R. Cahoon, H. W. Paxton

The possibility of using unidirectionally solidified, two-phase alloys as an approximation to fiber composite materials is investigated. The short-term me.chanical properties and failure modes of unidirectionully solidified A1 (rich)-Cu alloys containing ap -Proximately 0, 17.5, and 27.7 vol pct of 0 phase &apos;fibers" are determined at temperatures from 25" to 500" and compared with those obtained for conventionul SAP alloys. In a previous publication,&apos; hereafter referred to as I, the possibility of understanding some of the room-temperature mechanical properties of unidirectionally solidified castings was explored. For Al(rich)-Cu and Al(rich)-Mg two-phase alloys over a substantial range of compositions, the yield and ultimate strengths and common ductility measures were very adequately predicted from the principles of fiber strengthening4 and the analysis of ductility outlined by Gurland and Plateau." The results obtained in I suggest the possibility of using unidirectionally solidified, two-phase alloys to simulate fiber composite materials where the inter-dendritic second phase or constituent acts as the reinforcing material. Recent attempts concerning the fabrication of fiber conlposites have concentrated on producing composites with a good bond between fiber and matrix and with very long fibers so that their maximum contribution to the strength of the composite may be realized. However, these objectives are difficult to attain in practice and present fabrication processes are either extremely laborious or costly.13 The slow, unidirectional solidification of eutectics has received considerable attention as a method for producing composite materials. 5,6 This method can fulfill both of the above objectives but it is currently laborious, expensive, and has the additional disadvantage that the volume fraction of reinforcing phase cannot be easily varied. On the other hand, unidirectionally solidified, two-phase alloys, also with a good bond between the phases, are relatively easy to make and the volume fraction of reinforcing "fibers" can be easily varied by changing the average composition of the alloy. The disadvantage of the cast alloys is that the mechanical effectiveness of the "elongated interdendritic reinforcements" (EIR)* may be reduced due to their rela- tively short lengths, the w factor in Eq. [2] of I. However, if the EIR have a high strength their contribution can be considerable. For composite materials containing discontinuous cylindrical fibers of various lengths the ultimate strength is given by1 where it is assumed that the composite fractures when the fibers fail. In Eq. [I], a, is the stress in the matrix just prior to failure of the composite, Vf is the total volume fraction of fiber reinforcing constituent, Vf(l+) is the volume fraction of fibers whose lengths exceed the critical length, I,, which is defined as the shortest length of fiber in which the stress can build up sufficiently to break the fiber. af is the fracture strength of the fiber material, w is a factor accounting for the discontinuity of those fibers whose lengths exceed I,, 1-/d is the average aspect ratio of those fibers whose lengths are shorter than I,, and t is the shear stress in the matrix at the fiber-matrix interface. The factor w is dependent on the length of the fibers and also on whether deformation of the matrix occurs plastically or elastically. However, for a given length of fiber, w is smaller when elastic deformation of the matrix is assumed.&apos; It is of interest to consider the properties of simple unidirectionally solidified, two-phase alloys at elevated temperatures in view of the possibility of using suitable modifications for high temperature service. Knowledge of the creep behavior of these materials is still rudimentary (although under active investigation) and the present paper concerns itself with short time tensile properties of some alloys similar to those investigated in I (i.e., unidirectionally solidified Al(rich)-Cu alloys). Unidirectionally solidified alloys containing 5.6, 17, and 23 wt pct Cu were tested parallel to the direction of solidification at temperatures from 25" to 500°C. In the present investigation, the alloys were homogenized for 2 days at 535°C giving a matrix of homogeneous a phase (5.2 wt pct Cu) and an interdendritic constituent (EIR) which was completely Q phase (53 wt pct Cu). EXPERIMENTAL Alloys of nominal composition 5.6, 17, and 23 wt pct Cu (containing approximately 0, 17.5, and 27.7 vol pct 8 phase, respectively, after homogenization at 535°C) were prepared by melting 1200 g of A1 (99.99 pct) in a high purity graphite crucible and adding the appropriate amount of freshly cleaned copper chips (99.9 pct). The molten alloy (at 700°C) was poured into a preheated graphite mold (also at 700°C) and the ingot unidirectionally solidified by impinging water on the steel baseplate of the mold. The alloy was degassed immediately

Jan 1, 1970

By David Griffiths

I am not an orator and have never said a word in public during my thirty years&apos; experience in the coal mines of Colorado; but, being requested to speak by our worthy chairman, and as I am deeply interested in the future welfare of this organization, 1 will make an effort to say a few words in behalf of the Rocky Mountain Coal-Mining Institute, which we are about to organize. As 1 understand it, it is the object and intent of this organization not merely to assemble occasionally for the purpose of discussing various problems in coal-mining, and thus to obtain the best possible results in economic operations, but its especial duty will be to devise ways and means whereby accidents in coal mines will be minimized. Figuratively speaking. we are about to plant a tree in the State House of Colorado, and desire its roots and branches to extend over Wyoming on the north, New Mexico on the south. and Utah on the west. So we expect this tree to bear fruit over a vast area. We must remember that the planting of a tree, even in good soil, is useless unless we cultivate, prune, and nurture its development. If we will not do so, the tree will not bear the desired fruit. So, in order to make the Rocky Mountain Institute a success, we must not only start it off in good shape, but we must keep on adding to its strength in various ways, so that it will become a known factor and a recognized power over this broad territory. In other coal-mining states and countries of the world much has been accomplished through similar organizations to reduce loss of life among men engaged in the production of coal, and there is no reason why our institute in the West should not be equally useful and attain the same results. There are many ways by which this organization can be made beneficial to the men engaged, in various capacities, in handling the "dusky diamond." I will endeavor to confine my brief remarks to the necessity of a crusade of an educational nature among three classes of men that are directly and indirectly connected with coal mines. First-We need to educate ouselves. It is amazing how few of our number, holding positions of trust, and having charge of men in and around coal mines, are conversant with our brief mining laws and cognizant of our duties in enforcing them. Going down the line to our company men and miners, I think 1 can safely say that not over 5 per cent of them are in any way familiar with the statutes that have been enacted for their especial benefit and protection. This condition, 1 think, can be, and ought to be, remedied, and. if thoroughly done, the effect in reducing loss of life will soon become apparent. Now, to reach the officers or men in charge I consider an easy matter, which could be accomplished in this manner: Have our general managers or owners of coal properties issue a circular, compelling every superintendent pit boss, fire boss, or other officer having charge of men, to send them a copy of the mining law in their own handwriting every six months. But to come to the masses engaged in coal-mining. to get these men of various nationalities educated in the mining law, will be a difficult task indeed; but I think I have an original scheme that is practical, and which, if put in operation, would work wonders in educating our company men and miners. This plan is as follows: Have every school in every coal-mining district devote one-half hour on Friday afternoon after school hours to the familiarizing of all male pupils between the ages of twelve and sixteen years with the contents of our laws governing coal mines. At the end of each term of school, award small prizes- say, $5, $3, and $2- to those that are best posted, the principal of the school to be the sole adjudicator. Now, someone may say that this is causing extra work to our teachers, with no compensation. In this respect, I will vouch that 95 percent of our teachers will do this work gratis, providing we can show them the benefit of the system. If we go a step farther, and put an incentive before our teachers, probably it would be a good scheme to award a prize-say, $20 or more-for the best set of practical amendments to our mining laws complied by any school and its teachers, judges of the amendment to be members of this Institute and appointed by the President. There is no denying the fact that we need to be educated, and that we should adopt some system that will encourage a knowledge of practical and scientific mining amoung our present forces and the rising generation. Newspaper Men Need Educating Second-Our reporters and press agents need to

Jan 1, 1987

By R. L. Bell, T. G. Langdon

A technique was developed- for determining the grain strain, and hence the grain boundary sliding contribution, occurring during the high- temperature creep of a magnesium alloy, from the distortion of a grid photographically printed on the specimen surface. The results were compared with those obtained from measurements of grain shape, both at the surface and interrwlly, and it was concluded that the grain shape technique may substantially underestimate the grain strain and overestimate the sliding contribution due to the tendency for migration to spheroidize the grains. ALTHOUGH a considerable volume of work has been published on the role of grain boundary sliding in high-temperature creep, many of the estimates of Egb (the contribution of grain boundary sliding to the total strain) have been in error due to the use of incorrect formulas or inadequate averaging procedures.&apos; One of the most easy and convenient measurements from which to compute Egb is that of v, the step normal to the surface where a grain boundary is incident. Unfortunately, this parameter is also the one associated with the treatest number of pitfalls. Values of v have been used to calculate Egb from the equation: egb =knrVr [1] where k is a geometrical averaging factor, n is the number of grains per unit length before deformation, v is the average value of v, and the subscript ,r denotes the procedure of averaging along a number of randomly directed lines. If the dependence of sliding on stress were assumed, it would be possible, in principle, to calculate k from the known distribution of angles between boundaries and the surface. This in itself is difficult because the distribution depends on the history of the surface,&apos; but the problem is even further complicated by the fact that v depends on other factors such as the unbalanced pressure from subsurface grains.3 However, the great simplicity of the measurement procedure for v makes it highly desirable that this problem of k determination should be overcome. In the present experiments, this was achieved by the use of an indirect empirical method in which the grain strain, eg, at the surface was determined by the use of a photographically printed grid. The assumption here is that the total strain, et, is simply the sum of that due to grain boundary sliding, egb, and that due to slip or other processes within the grains, eg. SO that: Et = Eg + Egb [2] Thus k is given by: In practice, it is customary to indicate the importance of sliding by expressing it as a percentage of the total creep strain; this quantity is termed y (= 100Egb/Et). The determination of Eg from a printed grid within the grains avoids the difficulties due to boundary migration which should be considered when the grain strain is calculated from measurements of the average grain shape before and after deformation. As first pointed out by Rachinger,4,5 however, this latter technique has the particular advantage that it can also be applied in the interior of a polycrystal. Recently, several workers have produced evidence on a variety of materials6-&apos;&apos; to support the observation, first made by Rachinger on aluminum,4,5 that 7 can be very high, 70 to 100 pct, in the interior, even when the surface value, determined from boundary offsets, is very much lower.10&apos;11 Although there have been criticisms both of the shortcomings of the grain shape technique&apos;&apos; and of the different procedures used to determine y at the surface,&apos; it seemed important to check whether measurements of sliding by grain shape gave values of y which were truly representative of the material. In the present experiments, grain shape measurements were therefore made both at the surface and in the interior for comparison with one another and with the independent measurements of grain strain using the surface grid technique. EXPERIMENTAL TECHNIQUES The material used in this investigation was Magnox AL80, a Mg-0.78 wt pct A1 alloy supplied by Magnesium Elektron Ltd., Manchester. Tensile specimens, about 7 cm in length, were prepared from a 1.27-cm-diam rod, with two parallel longitudinal flat faces each approximately 3 cm in length. The specimens were annealed for 2 hr in an oxygen-free capsule, at temperatures in the range 430° to 540°C, to give varying grain sizes, and, prior to testing, the grain size of each was carefully determined using the linear intercept method. This revealed that the grains were elongated -0.5 to 5 pct in the longitudinal direction. Testing was carried out in Dennison Model T47E machines under constant load at temperatures in the range 150" to 300°C. At temperatures of 200°C and below, tests were conducted in air with the polished flat faces coated with a thin film of silicone oil to prevent oxidation; at higher temperatures, an argon atmosphere was used. To determine v,, each test was interrupted at regular increments of strain and the specimen removed from the machine. At the lower strains, when v, was less than about 1 pm, measurements were taken on a Zeiss Linnik interference microscope;

Jan 1, 1969

By W. A. Backofen, R. L. Fleischer

Single crystal, bicrystal, and polycrystal tensile tests of aluminum at 4.2°K, 77°K, and 300°K have been used to examine the role of grain boundaries in the deformation process. Results indicate that a grain boundary may affect the extent and slope of easy glide. The stage II hardening rate, on the other hand, is independent of the presence or absence of grain boundaries. This conclusion allows the size of the region of multiple slip caused by an incompatible grain boundary to be determined. For the size of bicyystal sample used in this study, multiple slip occurs in about half of the cross section. PREVIOUS studies of the stress-strain characteristics of bicrystals of face-centered-cubic metals have been limited to aluminuml-5 at room temperature. Recent results, however, indicate that the stress-strain curves of single crystals of such metals may be separated into at least three stages6 in which different deformation processes are occurring7 provided testing is done at sufficiently low temperatures.&apos; Since for aluminum a well-defined stage II develops only below room temperature, previous studies have not been able to relate effects of grain boundaries to all of the three stages of deformation. It is therefore to be expected that low-temperature deformation of aluminum single crystals and bicrystals should clarify the effects of grain boundaries on the different processes of deformation. EXPERIMENTAL PROCEDURE Single crystals and bicrystals were grown from the melt by the standard techniqueg with aluminum reported by Alcoa to be 99.993 pct pure. Ridges in the boat were used to guide the grain boundary during growth, assuring that the boundary would bisect the sample.10 The rate of furnace motion during growth was 1.0 cm per hr. During growth zone purification resulted, as evidenced by the ability of the first material to freeze to recrystallize at room temperature following severe deformation. Samples were approximately 4.4 X 6.6 mm in cross section and 103.5 mm in length between grips. Samples were annealed at 635" i 5°C for 40 hr and furnace cooled over a 7-hr period. They were then electropolished in a solution of 5 parts methanol to 1 part perchloric acid at a current density of 15 amp per sq dm for about 30 min at temperatures below 0°C. Tensile testing was performed at 295" (room temperature), 77" (sample in liquid nitrogen), and 4.2"K (sample in liquid helium) on the hard-type machine indicated schematically in Fig. 1. The machine con- sists basically of a tube surrounding a rod; one end of the sample is attached to each member, and the rod is pulled up the tube to extend the sample. The rod is rigidly mounted and is moved vertically by a system described by asinski." The pulling force is measured continuously by an electrical strain gage load cell, and the relative displacement of the tube and rod is also recorded continuously by a soft cantilever beam with electrical strain gages. Maximum stress and strain sensitivities were ±2g per sq mm and * 3-10-5. In all tests the strain rate was approximately 5.10-5 per sec. The thin wires in the tensile apparatus introduce softness, which may be corrected for, however, by measuring load vs displacement with the sample replaced by an elastic member. For loads greater than 15 kg the spring constant is 1.875.106 g per cm. The flexible wires also served to reduce substantially the large shearing forces which may arise in the case of grips having horizontal rigidity.&apos;&apos; As in any gripping system, however, bending moments will arise in the course of deformation by single slip. Engineering stress, s = (load)/(original cross-sectional area), and strain, E = (increase in length)/ (original length), are used for stress-strain curves unless otherwise indicated. Tables list resolved shear stress, T=mo and shear strain ? = dm, where m is the usual Schmid resolved shear stress factor for the primary slip system at the start of deformation. The first group of samples to be described forms an isoaxial set, all of the crystals making up the single crystals or bicrystals having the same tensile axis, the orientation of which is indicated by the cross in Fig. 2. For this orientation the primary slip plane and slip direction make angles of 45 deg with the tensile axis and the Schmid factor m has its maximum possible value of 0.5. Rotations about the tensile axis are indexed by means of an angle 0 between the small-area surface of the samples and the projection of the primary slip direction onto the cross section, as defined in Fig. 3. In single crystals, values of 0 were 0 and 90 deg, while in bicrystals 0 values were (0 deg, 180 deg), (90 deg, 270 deg), and (0 deg, 90 deg) as indicated in Fig. 4.

Jan 1, 1961

By R. R. M. Cousens, W. S. Garrett

Discussion: L. T. Campbell Pitt: When Mr Garrett and I discussed the design and construction of plugs and bulk head doors to resist great water pressures in mines about ten years ago, we visualized an emeregeny when there would be no time for site preparation. We could not however, anticipate a problem of such magnitude and urgency as the one described in the paper under discussion. The paper very clearly describes the flooding and the measures taken to save the mine. The complete flooding would have been a national disaster. The account of the combat between man and water reads like the account of a military operation. It was an endeavour which achieved success because there was a master plan, tenacity and, in fact, heroism. The paper is of necessity a long one. In my contribution I will mention some historical facts and factors that I believe have relevance to the flooding and the saving of the mine. It is of great satisfaction to me to have been one of the two initiators of a new approach to underground plug design and construction. The investigation was taken over by a Chamber of Mines Committee which combined our early work with further research and the resulting data had a significant bearing on the West Drie emergency plug designs. The final plug length was correctly based on the requirements to obviate leakage. A plug to have structural strength only would have been very much shorter but so ineffective as to have become a waste of labour, material, and in consequence, wasted the limited time available to prevent flooding of the whole mine. The figure of 75 lb per sq in. sheer stress provided a very ample margin, in my opinion, but under the very difficult site conditions and the limitations of plug material suitable for the very unusual construction, justified such a figure. There must have been considerable temptation to reduce the safety margin due to the urgency to complete and commission the plugs, but it is much to the credit of the directors of this operation that nothing less than a safe design was permitted. In about 1945 Blyvooruitzicht Gold Mining Company, which was then mining in to the West Driefontein area to assist in early development, decided to protect themselves from a possible inrush of water from the latter mine. A bulk head door was accordingly designed and installed on 2 level between 1945 and 1948 at 4,218 ft below surface. Since West Driefontein paid for it, Gold Fields consulting mechanical engineers collaborated with their colleagues of Rand Mines who completed their design after a few modifications were agreed. This door was never used to prevent an inrush into Blyvooruitzicht but the exercise was very useful to Gold Fields when the high pressure doors initially installed on 10 and 12 levels between Nos. 2 and 3 shafts to enable isolation of one section of the mine from another were designed. At that time the capacity of West Driefontein to cope with an inrush of water was comparatively small because the storage in stoped areas and the installed pump capacity were much less than in later years. These doors together with others of the same design were later moved to the lower levels served by sub vertical shafts. In addition, doors to prevent flood water entering the pump chambers on 18 level were installed at each end of that level, and designed for a head of 360 feet so that water would have to reach 16 level before the pumps were flooded. All these doors were called upon to hold water and it is reported that only a small plug in one of the doors and a 4 in. dia flanged pipe joint failed. The lesson is that in these cases the smallest details warrant as much attention as the major ones. That these were the only equipment failures when so much depended on equipment, is very much to the credit of design engineers, both of the Gold Fields head office and companies who supplied both mechanical and electrical equipment. Until I had read the paper I wondered why the policy of isolation by bulk head doors was not continued as the development moved eastward towards the Bank Compartment and particularly when it pierced the Bank dyke. The reasons have been given in the paper. It was unreasonable to expect provisions to be made to counter the remote contingency that the large installed pumping capacity and mine capacity to store water would be exceeded during a peak inrush. When pumps were being considered for West Driefontein, it was decided that the Gold Fields standard underground main pumps at that time of 1 million gallons per 24 hours would be too small. Three alternatives were considered. 1. Very large units that would require underground sub assembly and dismantling of both pump and motor. 2. The largest pump and motor that could be conveyed as a complete pump or motor from surface to the pump chamber. 3. Pumps of higher speed to reduce dimensions and weight. It was decided to adopt alternate 2 because of avoidance of transporting pumps in parts and assembling underground. Alternative 3 was not favoured because high speeds would have introduced problems of more exact rotor balancing. I believe the decision was correct and in the crisis enabled comparative easy transport of pumps and motors to surface for overhaul and, if required, transfer to other pump stations. The standard pump thus became a 10 stage 1,500 rpm pump with a capacity of approximately 2 million gallons per 24 hours against a static head that varied

Jan 3, 1969

By M. I. Brittan

Discussion Dr R. E. Robinson (Fellow): The author must be congratulated on a very meticulous and self-contained piece of work. It is indeed a pleasure to read a paper that is so clearly and systematically laid out, and where the conclusions and the testwork conducted have been so clearly described. The paper is complete in itself, which makes it very difficult for someone who is not intimately involved in the whole Torco project to make any comments on its content. However, two points that, strictly speaking, fall outside the scope of the paper are of considerable interest. The first relates to the particle size of the material treated. In the paper, the testwork is confined to one standard particle size (minus 60 plus 100 mesh). The essential feature of the paper is to indicate that the rate-controlling reaction in the whole segregation process is the rate of reaction of the ore particles with the reducing agent and the hydrogen chloride. One wonders, therefore, to what extent this relatively slow rate of reaction is affected by the particle size of the ore itself. One imagines that the reaction must take place by contact of the hydrogen chloride with the surface of the mineral particles, and it is reasonable to suspect that the rate of diffusion of the copper ions to the surface is a relatively slow process and is thus the limiting factor in this particular rate of reaction. It is possible, for example, that the improvement obtained, when the ore is subjected to reducing conditions before the chlorination, is due to a breakdown in the crystal structure of the original particle. This breakdown is brought about by the reduction and by the consequent increase in surface area available for reaction with hydrogen chloride. Can the author indicate whether any work has been done along these lines, and whether it has been established that the reaction depends on the surface area available? The second point relates to the application of this kinetic study to the actual operation of a Torco reactor. It was once planned to feed the sodium chloride, together with the reducing agent, into the top of the segregation chamber. In the paper, the author mentions that it has now been established that the segregation chamber behaves, to all intents and purposes, as a fluidized bed, and that there is, therefore, a rapid evolution of gas in the lower regions of the chamber, which, it is imagined, displaces the gas phase rapidly. Since the reaction between sodium chloride, water vapour, and the aluminium silicates in the ore is extremely rapid, one wonders how much of the hydrogen chloride produced is removed from the reaction zone before it has had time to react with the copper minerals. The extremely low consumption of sodium chloride (which is a vital feature of the Torco process) must depend on an extremely rapid circulation of the hydrogen chloride gas to all the ore particles in the segregation chamber. One wonders, therefore, if a system for the introduction of the sodium chloride into the bottom regions of the chamber might not result in even greater efficiency in the utilization of sodium chloride. ProC D. D. Howat (Fellow): All of us who have been concerned with the study of chemical reactions at high temperatures are keenly interested in kinetics and are well aware that this is not an easy study experimentally. Dr. Brittan is to be congratulated on the development of neat experimental methods and for his full discussion of the results obtained. Although the segregation process for the extraction of copper from oxide and silicate ores has been known for almost fifty years, the fundamental chemical and physical changes involved have been little understood and the fundamental data are very scanty. The work now in progress at A.A.R.L., together with that sponsored by the Anglo American Corporation in other research institutions throughout the world, is bound to produce new fundamental data and a much more complete understanding of this rather fascinating process. It is already apparent that some of the old and well-worn chemical reactions that were postulated to occur, just cannot take place in the way which was formerly accepted. Dr Brittan&apos;s work, carefully conducted and thoroughly analyzed as it has been, still leaves us with one great outstanding problem. The thermodynamic data and the possible reactions set out in Table I (page 281) of his paper leave us asking, in complete despair, how can copper be converted into a volatile chloride in the presence of HCl, CO and carbon at temperatures about 800°C? The thermodynamics all combine to show that copper should be reduced to metal as the first step in the process. This brings us right up against the second problem. If copper were reduced to the metal how would HCI convert it to the volatile chloride? On top of these problems is the unknown reason for the very high speed of reaction between CO, HCI and the ground copper ore. Still further into the region of the unknown is the reaction by which gaseous hydrochloric acid is produced in the actual process. Perhaps Dr Brittan is feeling grateful that he doesn&apos;t have to try to explain this reaction-at this stage of the research programme at least. The results very clearly show that both CO and HCl gas are essential for rapid production of the volatile copper chloride. Dr Brittan states that 18 minutes were required to attain 83 per cent extraction with HCl gas alone and this was reduced to 4 minutes when CO was

Jan 2, 1970

By W. A. Backofen, M. L. Ebner

SINCE the early work of Gough with Hanson and Wright,l-3 the study of fatigue has been characterized by experiments on single crystals only in recent times.9-10 Now, increasing attention is given to this aspect of fatigue research for the insight that it may provide into details of mechanism. The investigations have concentrated to a large extent on the development of deformation markings on fatigued crystals, and have shown the cracks to originate in slip bands possibly preceded or accompanied by slip-band extrusions. Experiments of special interest to the present work were conducted by paterson5 on copper crystals and involved both metallographic examination and measurement of change in flow stress. Crystals were cycled in alternating tens ion-compress ion with a constant plastic shear-strain amplitude of approximately 0.8 pct, and were particularly revealing for their demonstration of hardening with accumulated strain similar to that in unidirectional straining, through an easy-glide stage I followed by a stage II of rapid hardening; deformation was not continued beyond 40 cycles, however, so that the eventual course of the hardening curve could not be decided. For the conditions used by Paterson, surface slip markings were similar to those observed in unidirectional straining, but there were no X-ray asterisms and no deformation bands on the surface. In current thinking about the nature of fracture in fatigue, two views relative to mechanism are generally acknowledged, with the reservation that both could apply simultaneously in some measure. As one possibility, fracturing across a slip plane is regarded as a result of loss of cohesion from the creation of many point defects by dislocation movement under the cyclic loading.&apos;&apos; On the other hand, fracture has also been taken to follow as a consequence only of the geometry of slip at a free surface, consisting of offsets and crevices which eventually become fatigue cracks.12, l3 The work of McCammon and Rosenbergl4 showing fatigue in polycrystals at 4.2oK makes clear that any long-range diffusion of point defects is unnecessary, yet studies such as those of Forsyth and stubington15 show that accelerated diffusion may be a characteristic of deformation under fatigue loading. Interest in obtaining data of possible use for resolving such questions led to the experiments described below. Copper crystals were conveniently loaded in alternating four-point bending at constant deflection, a test condition shown to approximate a constant plastic strain amplitude for a wide range of axial orientations. The method of testing was not readily adapted to extensive study of temperature effects, but an investigation of the geometry of crack formation could simply be made by orienting,the slip direction at different angles to the surface. EXPERIMENTAL PROCEDURES Single crystals were grown by a modified Bridgman method in a stationary gradient furnace under an atmosphere of purified dry nitrogen. Purity of as-grown crystals was 99.999 pct as determined by spectrographic analysis plus vacuum fusion and gravimetric analyses for oxygen and sulfur, respectively. Crystals were of reasonable perfection as evidenced by a critical resolved shear stress at 10 deg from (110) of 60 g per sq mm and half-width of a (400) X-ray diffraction line from one crystal of

Jan 1, 1960

By R. R. Estill

THE rank of the Ruhr coal ranges from a high volatile bituminous coal to an anthracite, depending to some extent on the original depth of the seam. The average Ruhr coal corresponds to a soft bituminous American coal of a coking quality. The average thicknesses of individual coal seams being mined are also comparable (59 in. against 65 in. in the United States). However, consideration of seam conditions and mining conditions other than those just mentioned emphasizes differences rather than similarities with United States soft coal. In general, the Ruhr seams now being mined are much more folded and inclined than American seams. Dips of 20&apos; and 30" are common in seams now being worked, and 30 pct of the coal reserves in the district are in seams dipping more than 35". Only on the tops and bottoms of folds do we find rather flat coal seams. In addition to the folding there is extensive displacement by cross faulting plus a certain amount of strike faulting of an overthrust nature, which results locally in doubling or omission of seams. Because of the long history of mining in the Ruhr, nearly all coal lying near the surface has long since been mined out, and we find that the average depth of mining is at present about 2300 ft below the surface. Deep mining, folding, and faulting result in seam conditions requiring a great deal more roof support than one finds in American soft coal mines. In fact only in the anthracite district and the Rocky Mountain and Pacific coal fields do we find somewhat similar conditions. It is easy to say, therefore, that the problem of mechanization of coal cutting and loading in the German mines is quite different from that which we have so effectively met in America with our mobile cutters and loaders, duck bill loaders, and a room and pillar system of mining our drift and slope mines. Partly because of more limited coal reserves, the traditional German mining system is largely the longwall method, which gives an almost complete coal recovery. Backfilling must be extensively practiced to protect the longwall faces, the over and underlying seams and workings, and especially the surface industrialized areas and barge canals. The German engineers have accordingly concentrated their efforts on the design of cutters, loaders, and conveyors suitable to longwall methods rather than room and pillar methods. Undercutters with cutter bars like American models have been in use in the Ruhr since well before World War 11. In 1941 they accounted for 8.5 pct of the production. This percentage, of course, includes coal which was undercut but nevertheless had to be broken down with air hammers or with explosives. The most common of these cutters is the Eickhoff Standard cutter (see fig. 1). This machine does about 95 pct of the undercutting in the Ruhr today, and is available with either compressed air or electrical power and in at least four different sizes. A variation of the cutter is this one with two cutter bars (fig. 2). At the end of 1947 about 200 of these machines and similar cutters were accounting for 13.2 pct of the total production, a production which was, however, only 60 pct of the 1941 production rate, so that the actual cutter tonnage was only up to a small amount over 1941. In 1941 about 3 pct of the production was accounted for by shearing machines making their cut perpendicular to the longwall face. They were similar to those used in the States. These machines are today considered obsolete and now account for only 0.7 pct of the total production. They are located at only a few mines and at present do not seem to have much of a future in the Ruhr. For the future, the Ruhr miner is looking forward to rather extensive mechanization of face work, with two major types of equipment being developed almost simultaneously. On one hand there is the development of cutter loaders for use in relatively hard coal. They represent the further extension of ideas developed after relatively long experience with the Eickhoff cutter. On the other hand there has been since 1942 an intense interest in the Ruhr in the development of face-stripping methods, particularly by the Kohlenhobel (coal plow) and its modification. At the end of 1947 these cutter loaders, Kohlen-hobels and scrapers together were actually accounting for only about 1.4 pct of total production while air hammers still broke 77.1 pct and as much as 1.2 pct was actually broken by hand picks. However,

Jan 1, 1951

By Stephen B. Castor

The rare earth elements (REE) which include the 15 lanthanide elements (Z = 57 through 71) and yttrium (Z = 39) are so called because the elements were originally isolated in the late 18th and early 19th centuries as oxides from rare minerals. Most REE are not as uncommon in nature as the name implies. Cerium, the most abundant REE (Table 1), comprises more of the earth&apos;s crust than copper or lead. Many REE are more common than tin and molyb¬denum, and all but promethium are more common than silver or mercury (Taylor, 1964). Promethium (Z = 61) is best known as an artificial element, but has been reported in very minute quantities in natural materials. Lanthanide elements with low atomic numbers are generally more abundant in the earth&apos;s crust than those with high atomic numbers. In addition, lanthanide elements with even atomic numbers are two to seven times more abundant than adjacent lan¬thanides (Table 1) with odd atomic numbers. The lanthanide elements traditionally have been divided into two groups: the light rare earths (LREE), lanthanum through eu¬ropium (Z = 57 through 63); and the heavy rare earths (HREE), gadolinium through lutetium (Z = 64 through 71). Although yttrium is the lightest REE, it is usually grouped with the HREE to which it is chemically and physically similar. The REE are lithophile elements (elements enriched in the earth&apos;s crust) that invariably occur together naturally because all are trivalent (except for Ce+4 and Eu+2 in some environments) and have similar ionic radii. Increase in atomic number in the lanthanide group is accompanied by addition of electrons to an inner level rather than the outer shell. Consequently, there is no change in valence with change in atomic number, and the lanthanide elements all fall into the same cell of the periodic table. The chemical and physical differences that do exist within the REE group are caused by small differences in ionic radius, and generally result in segre¬gation of REE into deposits enriched in either light lanthanides or heavy lanthanides plus yttrium. The relative abundance of individual lanthanide elements has been found useful in the modelling of rock-forming processes. Comparisons are generally made using a logarithmic plot of lanthanide abundances normalized to abundances in chondritic (stony) meteorites. The use of this method eliminates the abundance vari¬ation between lanthanides of odd and even atomic number, and allows determination of the extent of fractionation between the lanthanides because such fractionation is not considered to have taken place during chondrite formation. The method is also useful because chondrites are thought to be compositionally similar to the original earth&apos;s mantle. Europium anomalies (positive or negative departures of europium from chondrite-normalized plots) have been found to be particularly effective for petrogenetic modelling. REE were originally produced in minor amounts from small deposits in granite pegmatite, the geologic environment in which they were discovered. During the second half of the 19th century and the first half of the 20th century, REE came mainly from placer deposits. With the exception of the most abundant lanthanide el¬ements (cerium, lanthanum, and neodymium), individual REE were not commercially available until the 1940s. Since 1965, most of the world&apos;s REE have come from two hard rock deposits: Mountain Pass, United States, and Bayan Obo, China. GEOGRAPHIC DISTRIBUTION OF REE DEPOSITS More than 70% of the world&apos;s REE raw materials come from three countries: China, the United States, and Australia. China emerged as a major producer of REE raw materials during the 1980s, while Australian and United States market share decreased dramatically (Fig. 1). Table 2 gives recent annual production figures along with estimated reserves by country, and Fig. 2 shows loca¬tions of significant REE mining. MINERALS THAT CONTAIN REE Although REE comprise significant amounts of many minerals, almost all production has come from less than ten minerals. Table 3 lists minerals that have yielded REE commercially or have po¬tential for production in the future. Extraction from a potentially economic REE resource is strongly dependant on its REE miner¬alogy. Minerals that are easily broken down, such as bastnasite, are more desirable than those that are difficult to dissociate, such as allanite. In general, producing deposits contain REE-bearing min¬erals that are relatively easy to concentrate because of coarse grain size or other attributes. For more thorough discussions of REE¬bearing minerals see Mariano (1989a) and Cesbron (1989).

Jan 1, 1994