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By T. Saburi, C. M. Wayman

The massive and martensitic transformations in ß Cu-Ga alloys were studied by optical microscopy and by transmission electron microscopy and diffraction. These types of transformation are distinct and do not merge into one another. An internal fine structure due to stacking faults was found in both the massive phase (m) and the martensite, but the origin of the faults is believed to be different in each case. The faults in the martensite are due to the lattice invariant or "second" shear of the phenom-enological theory, while those in the are probably growth faults and deformation faults arising from the transformation volume change. Evidence is presented for ordering in the ß phase prior to transformation into martensite, and the crystal structures of and ß' and ßi martensites are discussed. THE results from some previous investigations1,2 have led to the suggestion that martensitic and massive transformations have certain common characteristics. It presently seems established that massive transformations like martensitic transformations are essentially diffusionless,1,2 but unlike martensitic transformations, when a new phase forms massively there is no shape change or surface relief.= If the massive-type transformation is crystallographically similar to martensitic transformations, one would expect a fine structure as a result of the lattice invariant shear4 which is a part of the transformation mechanism. The existence of such a fine structure (i.e., slip, twinning, or faulting), if any, can be verified by transmission electron microscopy, and thus the present investigation was undertaken to compare the internal structure of massively and martensitically formed phases in Cu-Ga alloys which were derived from a parent of the same composition. The transformation characteristics of ß Cu-Ga alloys were also studied. The bcc ß phase in the Cu-Ga system, Fig. 1, is stable above 616oc.' Depending upon the cooling rate, ß Cu-Ga alloys can undergo three different types of phase transformations. The ß phase of eutectoid composition (23.7 at. pct Ga) decomposes into the ? phase (hcp) and the phase (complex cubic) by means of a eutectoid reaction. When the cooling rate is increased so that the decomposition of 0 into [ and ? is suppressed, the ß phase transforms into a nonequilibrium hcp phase, herein denoted ?m by a process of very rapid growth called a "massive" transformation. At even faster quenching rates the massive transformation can be suppressed and the ß transforms martensitically. Massive transformations, apparently similar to those in the Cu-Ga system, have also been studied in the Cu-Zn2,6-8 and Ag-Cd systems.9 Definitive characteristics if the Massive Transformations in 0 Cu-Ga Alloys1 1) The most remarkable feature of the massive transformation in Cu-Ga is that there is no bulk composition change on transformation. According to Massalski,1 the resulting [, phase is a supersaturated extension of the equilibrium [ phase. The lattice parameters continue the trend exhibited

Jan 1, 1965

By E. R. Borcherdt

METHODS and equipment are subject to study in the Butte district at all times in order that advantage may be taken of any new developments that might serve to increase efficiency and lower costs. Detailed time studies of each operation involved in the entire mining process have been of invaluable assistance in analyzing the individual operation for possible improvement. In other words, the fact that a practice has been in vogue for many years does not exempt it from change provided an easier and cheaper method can be found. EXPLORATION Ore occurs in the Butte district in such manner that structures must be drifted on, no matter how small or unimportant they appear when intersected by a crosscut. Parallel structures also add to the crosscut requirements. This prospecting work, which formerly was done entirely by drifting and crosscutting, is now being supplemented by diamond drilling, to aid in greatly decreasing the amount of waste. In long-shot prospecting, or in second or third class prospecting, where the chances of developing commercial ore are not very great, diamond drilling is of particular value. In prospecting where chances of encountering commercial ore are certain or where crosscuts can be used later for haulage or ventilation, drilling obviously is not applicable. Improvement in drilling machines and bits has contributed greatly to the present success of diamond drilling. STOPING For many years square-set filled sloping was the principal mining method used in the Butte district. Later timbered and untimbered rill stopes were used, also some horizontal cut-and-fill sloping where both ore and walls were particularly strong. Because of the amount of timber required for square-set stoping, and the relatively low production per scope unit, experiments with horizontal cut-and-fill scopes were begun in ore bodies of which many had fairly weak walls. It was felt that increase in the speed of the stoping operation, followed closely by filling, to avoid weight being produced by permitting ore and walls to remain insufficiently supported over comparatively long periods, would maintain the walls and backs in safe condition without the use of timber and without ore dilution from wall slough. Cycles of drilling, scraping ore, and filling were devised, which have materially increased stope production and decreased timber consumption, so that both stope labor and supply costs were lowered appreciably without materially affecting grade control. In the larger mines of the district, where wide veins favor this stoping method, half of the total production is now being mined by horizontal cut-and-fill stopes (Fig. I). The increase of the filling rate demanded by the increased stoping rate was the greatest obstacle encountered in the accelerated stoping operations. Finally this problem was solved by replacing the time-honored method of transferring filling waste by small cars, which were hoisted or lowered in the auxiliary shaft, by a transfer

Jan 1, 1941

By H. K. Work

THE Bessemer process is one of the most interesting methods of making steel. At one time it was by far the most important. In recent years, however, it has steadily lost ground to the open-hearth process. The forces working against this method of steelmaking are partly economic and partly metallurgical.' In this paper, dealing as it does with the control of the process, the metallurgical factors alone will be considered, although it is obvious that economic factors must play a part in determining the ultimate status of the Bessemer process. This discussion has been limited in another respect-it deals primarily with the acid Bessemer process in bottom-blown converters, although much of the information is applicable to basic and side-blown converters. The fact that open-hearth steel has superseded Bessemer steel to a large extent has tended to create an impression that Bessemer steel is inferior. As a matter of fact, Bessemer steel is actually superior to open-hearth for many applications.2 There are, however, definite metallurgical limitations to the use of Bessemer steel, resulting largely from two factors: (I) the physical properties conferred by nitrogen and phosphorus, and (2) the variability from blow to blow. These, of course, are not entirely independent, since some of the variability covered in the second item is undoubtedly due to nitrogen. But the variability is also caused by other factors. One of these is the relatively small size of the Bessemer heats. Because of this, a given shipment of Bessemer steel is more likely than a shipment of open-hearth steel to contain the product of several heats, each with its own distinctive and slightly different characteristics. Added to this is the fact that the refining of Bessemer steel proceeds so rapidly that the timing of the end point is limited to a matter of seconds, whereas the finishing period in the open hearth extends over a period of one or two hours and the timing of the end point may vary by several minutes without seriously affecting the product. There is no doubt that this has contributed to the variability of Bessemer steel and has led to the use of open-hearth steel in many optional applications. That Bessemer steel may vary is not necessarily objectionable, and it may even prove an advantage, providing a precise control is available to secure the desired qualities for any specific application. With this in mind, a comprehensive research program was undertaken, one phase of which was directed toward securing more accurate control of the blow. The results have been encouraging, and it is the purpose of this paper to describe a method in which modern scientific instruments, based on photocells, are currently being employed to assist the blower. THE CONTROL PROBLEM In this day of scientific control of technical processes, where the operating

Jan 1, 1941

By T. J. Woods

Tars paper deals with the origin, definition and prevention of scabs an semifinished rolled-steel product. Mold coatings, which are considered essential in scab prevention, were found to be effective only in relation to the ingot surface of semikilled and killed steels and not necessary for the rolled product. That is to say, the scabs on ingots caused by having been cast in uncoated molds are eliminated in the soaking and rolling operations, therefore are not present on the rolled product. This fact, which is contrary to the average opinion, was established after extensive and thorough experiments had proved the ordinary conception of the origin of scabs to be incorrect. This subject may be of little or no concern to steel plants producing small ingots for the manufacture of small billets and bars, except as it may be of interest in showing why their surface troubles with scabs do not exist on the product of heats with normal pouring practice. But to the producer of large ingots to be rolled into slabs for large plates used in shipbuilding, it is a very important subject. DEFINITION OF A SCAB The generally accepted opinion is that the individual scabs present on rolled products (as shown by Fig. I) are considered to have been individual splashes on the ingot surface. This opinion was found to be incorrect. Experiments showed that there were no individual scabs on the product that had appeared as individuals on the ingot surface. The single scabs seen are either pieces of a shell that is present near the bottom of all top-cast ingots, or a veneer that may be present on any part of an ingot improperly poured, as when the stream strikes the mold wall. This definition is based on a detailed study of scabs and their characteristics in which the following have been found true: Location.- On the product of heats poured without trouble, scabs, if present, are always on the product that corresponds to the bottom end of the ingot, and usually occur in clusters, This is true whether molds are coated or not. Shape.-There is no regularity as to shape, size or thickness. It is sometimes possible to separate a few pieces and place them together, like pieces of a jigsaw puzzle, indicating that at one time they had been united as one piece. If individual splashes were the origin of scabs, the foregoing would not be true. Instead, the splashes, which occur throughout the pouring of the ingots, would be present uniformly over the entire product and there would be considerable similarity in the shape and thickness of the scabs. EXPERIMENTAL WORK Our experimental work was conducted on 200-ton heats of semikilled plate steel tapped into an oval ladle 15 ft. high. All ingots were top-cast direct into 24 by 54 by 82-in. molds, straight-sided, set on fat stools. Two separate studies were made about one year apart, each study extending

Jan 1, 1943

By E. M. Tory, P. T. Shannon

An analysis of batch and continuous thickening in terms of the movement of planes of constant concentration (i.e. continuity waves) is presented. Use of solids flux as a primary variable greatly facilitates the analysis and clarifies the relationship between batch and continuous thickening. An "ideal" slurry is defined as one whose settling rate is a function only of the local solids concentration. An analysis of continuous thickening of this ideal slurry in an ideal thickener explains many aspects of the steady — and unsteady — state behavior of real slurries, such as the main features of a thickener's "solids profile" (including the break attributed to rake action), the sudden appearance at abnormally high feed rates of a zone of constant concentration and the velocity with which it rises, and the effect of detention time at constant weight of sludge per unit area. Where deviations exist, they are explained in terms of the batch settling behavior of real slurries. The purpose of this paper is to interpret continuous thickening phenomena in terms of batch settling behavior. Both types of thickening may be considered as the process of propagating concentration changes, due to the downward movement of the solids, from the bottom of the settling vessel upward through the slurry.' These concentration changes may be either finite or infinitesimal as in the case of the propagation of a continuous concentration gradient. The slurry itself may be moving downward as in the case of continuous thickening. The downward velocity of the slurry may be such that the concentration changes being propagated up from below remain stationary relative to the thickener walls. This is the case for steady-state thickening. It seems appropriate to analyze continuous thickening in terms of the movement of planes of constant concentration. This has already been done for batch settling."3 Although our main interest is in real slurries, there are certain aspects of both types of settling which can be deduced from a theoretical analysis of highly idealized systems. To this end, an "ideal slurry" is defined as one for which the dimensionless settling rate at any point is a function only of the local solids concentration.1,2 An "ideal thickener" is one in which there is no radial variation in either the solids concentration or in the downward component of slurry velocity. The underflow outlet is a point sink. It is to be expected that the behavior of non-ideal systems will deviate in some measure from theoretical results. However, the same characteristics which cause deviations in continuous thickening should be evident in batch settling, permitting a rational explanation of the former in terms of the latter. PREDICTED RESULTS FOR AN IDEAL SLURRY IN AN IDEAL CONTINUOUS THICKENER General Theory: In continuous thickening, the downward solids flux (mass/area-time) relative to the thickener, G, results from two different

Jan 1, 1967

By M. C. Dow

Belt conveyors generally are conceded to be the most economical method yet devised for the transportation of large quantities of bulk materials within plants. Belts are coming into greater use for transportation over long distances and are replacing other forms of long-haul conveyance. It is not the intent of the writer to comment on the design of belt conveyors, but to offer some suggestions for improvement in the operation of conveying belts, based on many years experience with belts in several stone-crushing plants. It is the contention of this writer that opportunities for improvement not fully realized by most operators exist in many plants and that a study of the individual conveyors can result in lower unit belt costs. General Reliability It is the fool-proof nature and general reliability of the belt conveyor that is responsible for its neglect. Even when abused, a belt will carry large tonnages and last a long time, often outlasting the personnel responsible for its care. Some operators look upon the purchase of a new belt as an expensive outlay every so many years, rather than a constant and steady cost amounting in belt wear alone to several dollars every operating day. This belt wear for all conveyors in one stone-crushing plant known by the writer costs 1 /12 cents per ton or $12 to $15 per hour, based on the present cost of belt replacements. The fact that these replacement costs are double those of prewar quality emphasizes the necessity of improving the conditions that make belts wear out. Some of those conditions are the responsibility of the conveyor designer, and if a serious original error has been made, this handicap cannot be overcome by the operator. If the belt is too narrow, too steep, or too fast for good loading, the operator finds that he has an inherently expensive belt. The other factors that decrease the life of a belt are more or less under his control. The two factors of most importance are the loading conditions and the alignment. Most of the cover wear and carcass cutting occurs at the loading point and if the alignment is not good under all conditions, edge wear or damage may result, Fig I. Loading the Belt It is difficult to load a belt in an ideal manner; that is, to flow the material on the moving belt at a steady rate, without impact, without turbulence, and with a velocity equal to and in the same direction as the belt. An inclined belt is especially difficult to load and particularly when its speed is over 300 or 350 fpm. Inclined conveyors require longer skirt plates to prevent spillage until the load settles. It does not seem reasonable to expect an inclined the to carry the same load as a horizontal belt of the same width and speed, because of the difficulty in loading

Jan 1, 1948

By M. C. Dow

Belt conveyors generally are conceded to be the most economical method yet devised for the transportation of large quantities of bulk materials within plants. Belts are coming into greater use for transportation over long distances and are replacing other forms of long-haul conveyance. It is not the intent of the writer to comment on the design of belt conveyors, but to offer some suggestions for improvement in the operation of conveying belts, based on many years experience with belts in several stone-crushing plants. It is the contention of this writer that opportunities for improvement not fully realized by most operators exist in many plants and that a study of the individual conveyors can result in lower unit belt costs. General Reliability It is the fool-proof nature and general reliability of the belt conveyor that is responsible for its neglect. Even when abused, a belt will carry large tonnages and last a long time, often outlasting the personnel responsible for its care. Some operators look upon the purchase of a new belt as an expensive outlay every so many years, rather than a constant and steady cost amounting in belt wear alone to several dollars every operating day. This belt wear for all conveyors in one stone-crushing plant known by the writer costs 1 /12 cents per ton or $12 to $15 per hour, based on the present cost of belt replacements. The fact that these replacement costs are double those of prewar quality emphasizes the necessity of improving the conditions that make belts wear out. Some of those conditions are the responsibility of the conveyor designer, and if a serious original error has been made, this handicap cannot be overcome by the operator. If the belt is too narrow, too steep, or too fast for good loading, the operator finds that he has an inherently expensive belt. The other factors that decrease the life of a belt are more or less under his control. The two factors of most importance are the loading conditions and the alignment. Most of the cover wear and carcass cutting occurs at the loading point and if the alignment is not good under all conditions, edge wear or damage may result, Fig I. Loading the Belt It is difficult to load a belt in an ideal manner; that is, to flow the material on the moving belt at a steady rate, without impact, without turbulence, and with a velocity equal to and in the same direction as the belt. An inclined belt is especially difficult to load and particularly when its speed is over 300 or 350 fpm. Inclined conveyors require longer skirt plates to prevent spillage until the load settles. It does not seem reasonable to expect an inclined the to carry the same load as a horizontal belt of the same width and speed, because of the difficulty in loading

Jan 1, 1948

By T. E. Tietz, J. E. Dorn

Activation energies for creep of copper at intermediate temperatures, where crystal recovery was negligible, were determined by the simple technique of rapidly alternating the test temperature between and T2 (T2 = TI + about 10°K) throughout a constant stress creep test. The activation energy for creep H was found to be 37,000± 3,000 cal per mol, independent of stress and strain. The same creep laws as have been previously established for high temperature creep were found to be valid for creep at intermediate temperatures. But the AH was found to be lower than that for self-diffusion in the intermediate temperature range whereas it is known to be equal to that for self-diffusion at high temperatures. IT is now well established that creep at high temperatures can be correlated by the functional relationship'-" e = f(?) s = constant [I] where E is the total plastic strain, f is the function that depends on the stress, ? = f e -- dt, t is the duration of test, e is the base of natural logarithms, AH is the activation energy for creep, R is the gas constant,T is the absolute temperature, and s is the stress. Thus, identical values of are achieved during creep under a given stress at two alternate temperatures at the same total strain. At the same strain therefore where subscripts refer to the two alternate temperature conditions of test. Consequently, the activation energy for high temperature creep is readily obtained with the aid of Eq. 2. Activation energies so obtained have been observed to be insensitive to the creep strain at which they are evaluated, the applied stress, grain size, minor alloying additions,'-' cold worked state, and the presence of thermally stable dispersions of intermetallic compounds in an a solid solution matrix Furthermore, the activation energy for high temperature creep is very nearly equal to that for self-diffusion in all cases where adequate data are available for comparison." " The apparent identity of the activation energies for high temperature creep and self-diffusion strongly suggest that high temperature creep arises from a dislocation climb process. This hypothesis is further strengthened by the fact that the energy of activation for high temperature creep is insensitive to the applied stress," and by the observations that the correlation suggested by Eq. 1 cannot be ex- tended to temperatures below those for rapid crystal recovery. When using the activation energy for self-diffusion, Eq. 1 fails to give the required correlations between creep curves at various intermediate temperatures.' This suggests that the dislocation climb mechanism for creep might be superseded by some alternate mechanism in the intermediate range of temperatures. This investigation was initiated in a preliminary attempt to ascertain the type of creep law that might be valid in the intermediate range of temperatures. Special consideration, however, was given to the determination of the activation energy for creep in this range because the Becker-Orowan', " and the Mott-Nabarro10 theories for transient creep suggest that the activation energy should increase

Jan 1, 1957

By L. M. Barker, E. G. Lewis

THE literature of milling is replete with papers devoted to the subject of ball milling, all of which no doubt have contributed in one way or another to progress in that art. In this paper reference will be made to only a few of the many authors, but the value of the contributions of all who have written on that subject is hereby acknowledged. With a few exceptions the efforts of most students of the subject have been devoted to laboratory investigations of principles. Davis1 was one of the first to report on plant-scale investigations and to present to mill men certain criteria for the guidance of ball-milling practices. His conclusions with respect to the significance of sizing analyses of classifier sands in their relation to the factors of mill speed and ball sizing are particularly referred to. Those conclusions were: I. If the balls are large or the speed of the mill is high, crowding will appear at the finer sizes in the classifier sands. 2. If the balls are small or the speed is low, crowding will appear at the coarser sizes in the classifier sands. 3. The indications are that best efficiency is obtained when the screen analysis of the sands show a minimum crowding at any size. The last conclusion was qualified by the statement that it had not been proved conclusively. It is surprising, however, that the literature of later years reveals little evidence that millmen have made any serious attempt to apply those criteria. This may, perhaps, be explained by the fact that the average plant-scale ball-milling problem presents so many variables that an investigation of any one of them encounters the problem of control of the others-a problem that often is unsurmountable in a small plant. In a large plant, in which there is a duplication of grinding units in parallel operation, opportunity may be afforded for determining the value of such criteria. Nevertheless, a study of published data from grinding circuits of both large and small plants suggests that either Davis' criteria have been tried and found inapplicable or have never been given consideration. For the purpose of testing Davis' conclusions and with improvement of ball-milling efficiency as the final objective, a test program was initiated in the concentrator of the New Cornelia branch of the Phelps Dodge Corporation early in 1938. The practices of that plant have already been described .2 The fine-grinding units are arranged so that any one of eight units may be operated as a test unite and its results compared with those of the remaining units. All units receive similar feed from a common storage bin, in which the product of the fine-crushing plant is bedded by traveling tripper. The ball mills are all similar in size and operated at the same speeds, and are individually metered. The feeder conveyors of each unit are equipped with Merrick weightometers. The test program as initially outlined proposed investigation of the following:

Jan 1, 1941

By G. B. Walker, S. A. Falconer, S. J. Swainson, H. H. Vogel

During the past few years much interest and attention has been focused on a relatively new method of ore concentration, which utilizes the principles of sink and float and employs as the heavy medium a suspension in water of a finely ground solid, such as galena or ferrosilicon. Although a relatively recent development, much progress has already been made in applying heavy-media separation processes to a considerable number of ores. Eight plants in this country and abroad are now treating approximately 10 million tons of ore annually by heavy-media separation (sink-float) processes, including lead-zinc, lead, iron, garnet, tin and tungsten. At the annual meeting of the A.I.M.E. in 1941, three interesting papers (unpublished) dealing with heavy-media separation (sink-float) processes of ore concentration as practiced in four commercial units in the United States were presented by Victor Rakowsky, Grover Holt, Elmer Isern and Robert Ammon. These papers gave an excellent exposition of the principles of sink and float, as well as the operating practice in two iron-ore-beneficiation plants, the Harrison and the Merritt, of Butler Brothers, in Minnesota; the Central mill of the Eagle Picher Mining and Smelting Co.; and the Mascot mill, of the American Zinc, Lead and Smelting Co. Augmented by pictorial descriptions and a working model demonstrating the processes, these papers evoked a great deal of interest and discussion. In fact, in reporting the high lights of the meeting in the April 1941 issue of Mining and Metallurgy, Professor Taggart made the statement that " preconcentration by heavy suspension was Miss A.1.M.E 1941 so far as milling is concerned." Although to date the greatest application of heavy-media separation from the tonnage point of view has been for the treatment of lead, zinc and iron ores, these processes have a much wider field of usefulness. In this regard, it is to be noted that recently new heavy-media separation plants for treating cassiterite ore and garnet ore have been placed in successful operation. Moreover, numerous other types of ores have been found to respond well to these processes when tested on a semi-commercial scale, and it is expected that in the near future a considerable number of new plants will be erected. In view of the fact that heavy-media separation processes offer such attractive possibilities, particularly now, when the mining industry is being called on for an even larger production than ever before has been attempted, despite shortages of labor and supplies, it was felt that a description of some of the successful separations that have been worked out during the past year in the Ore Dressing Laboratory of the American Cyanamid Co. on a variety of ores of strategic importance would be of interest. The purpose of this paper, therefore, is to present the essential data relating to investigations of heavy-

Jan 1, 1943

By G. B. Walker, S. J. Swainson, S. A. Falconer, H. H. Vogel

During the past few years much interest and attention has been focused on a relatively new method of ore concentration, which utilizes the principles of sink and float and employs as the heavy medium a suspension in water of a finely ground solid, such as galena or ferrosilicon. Although a relatively recent development, much progress has already been made in applying heavy-media separation processes to a considerable number of ores. Eight plants in this country and abroad are now treating approximately 10 million tons of ore annually by heavy-media separation (sink-float) processes, including lead-zinc, lead, iron, garnet, tin and tungsten. At the annual meeting of the A.I.M.E. in 1941, three interesting papers (unpublished) dealing with heavy-media separation (sink-float) processes of ore concentration as practiced in four commercial units in the United States were presented by Victor Rakowsky, Grover Holt, Elmer Isern and Robert Ammon. These papers gave an excellent exposition of the principles of sink and float, as well as the operating practice in two iron-ore-beneficiation plants, the Harrison and the Merritt, of Butler Brothers, in Minnesota; the Central mill of the Eagle Picher Mining and Smelting Co.; and the Mascot mill, of the American Zinc, Lead and Smelting Co. Augmented by pictorial descriptions and a working model demonstrating the processes, these papers evoked a great deal of interest and discussion. In fact, in reporting the high lights of the meeting in the April 1941 issue of Mining and Metallurgy, Professor Taggart made the statement that " preconcentration by heavy suspension was Miss A.1.M.E 1941 so far as milling is concerned." Although to date the greatest application of heavy-media separation from the tonnage point of view has been for the treatment of lead, zinc and iron ores, these processes have a much wider field of usefulness. In this regard, it is to be noted that recently new heavy-media separation plants for treating cassiterite ore and garnet ore have been placed in successful operation. Moreover, numerous other types of ores have been found to respond well to these processes when tested on a semi-commercial scale, and it is expected that in the near future a considerable number of new plants will be erected. In view of the fact that heavy-media separation processes offer such attractive possibilities, particularly now, when the mining industry is being called on for an even larger production than ever before has been attempted, despite shortages of labor and supplies, it was felt that a description of some of the successful separations that have been worked out during the past year in the Ore Dressing Laboratory of the American Cyanamid Co. on a variety of ores of strategic importance would be of interest. The purpose of this paper, therefore, is to present the essential data relating to investigations of heavy-

Jan 1, 1943

By F. O. Garrabrant

Power is furnished to the Climax Molybdenum Co. by the Public Service Co. of Colorado over two 100,000-volt lines to a bank of three 3333-kva. transformers 100/13.8 kv. These transformers are so designed that with the addition of forced cooling their capacity may be increased 33? per cent at 12,000-ft. altitude. Under present conditions, these transformers are operated at 140 Per cent of their rating with forced air cooling. On the secondary side of this bank there is a 15-kv., 1200-amp. Westinghouse oil circuit breaker provided with automatic reclosing relays allowing three successive short circuits before lockout. From this breaker 13.8-kv. power is transmitted to a central switching station (Fig. I) where all circuits radiate to the different operations of the property. Each circuit is protected by a 23-kv., 600-amp. oil circuit breaker with by-pass and isolating disconnecting switches. Mine Power at 13.8 kv. enters the mine through a ventilation raise over a three-conductor, size 1-0 varnished cambric insulated, leaded, steel-wire armored cable, insulated from 17 kv. This power is received and distributed through one switching station and four underground substations, shown in Fig. 2. No. 370 substation (Fig. 3) contains three a 13,800/440-volt, single- phase Pyranol transformers protected on the high-voltage side by a 600-amp. manually operated oil circuit breaker. This station serves 20 slusher-hoist motors, a 45-kw., three-phase, 220-volt electric furnace, a Massco hot miller for tempering and milling Timken detachable bits and a hoist with 175-hp motor on the No, 3 development shaft. Also at this shaft are three pumps. One two-stage pump, 1200 gal, per min., 500-ft. head, 3600 r.p.m., driven by a 150-hp. motor, is at the bottom of the shaft, Two deep-well pumps, 500 gal. per min., 500-ft. head, are at the collar of the shaft. under normal operation the one pumping 1200 gal. per min, is controlled by a float switch to take care of the water. If this pump fails, or if its capacity is insufficient, the first 500-gal. pump is automatically started1 and if still more capacity is needed the second 500-gal. pump automatically starts. Time delay relays are provided on the deep-well pumps, so that if power fails the column will have time to drain before the Pumps restart. On failure of any Of these, an alarm is sounded in the train dispatcher's office, so that mechanics can be sent to ascertain the cause of failure and make repairs. The cable from No. 37' substation is run to the switching station, in which are four 15-kv., 600-amp. direct-current sole-noid-operated oil circuit breakers. The current for Operating and tripping is furnished by a 125-volt storage battery. All breakers are operated from the switchboard in the compressor room. One circuit from this switching station goes to the

Jan 1, 1946

By Ali Ebrahimi, Meena L. Viswanath, John F. Beech, Mustafa B. Erten, Ming Zhu

"Excessive deformation of a cantilever sheetpile wall can occur due to dredging, excavation, backfilling, or long-term creep behavior of soil. A buttressing berm constructed in front of the wall can be considered as a mitigation technique to prevent or decrease the deformation by increasing the passive earth pressure and thus improving the wall stability. The effect of the buttressing berm on the stability and deformation of a sheetpile depends on the berm geometry (i.e., the thickness/depth and the width) as well as the properties of the sheetpile and the adjacent soils. A parametric study was conducted using a conventional limit equilibrium method. Design charts were prepared to evaluate the effect of a buttressing berm and the soil friction angle on the wall stability and deformation. These charts can be used as a tool by geotechnical engineers to make a quick estimate of the required size of a buttressing berm in case of calculated or unexpected excessive deformation of a sheetpile. An analysis of the effect of a buttressing berm on the stresses and deformation of a sheetpile was also compared with the results from finite element analysis. The results of this comparison are discussed.INTRODUCTIONCantilever sheetpile walls are commonly used to support excavations. Sheetpiles are hot rolled or cold formed sections of steel made to interlock with adjacent sections to form a continuous structure providing lateral support. Sheetpile walls can be cantilever walls or anchored with tiebacks. To construct a cantilever sheetpile wall, sheetpiles are driven by pile drivers or vibrated in to a design depth sufficiently below the depth of excavation. Excavation can then proceed in front of the sheetpile wall, which resists the earth pressure loading from behind the wall. The design depth and geometry of the sheetpile wall depend on the soil conditions, the depth of excavation, and the tolerable deformation. In certain cases, when a small buttress of soil can be left immediately adjacent to the sheetpile wall, the required depth or geometry of the sheetpile wall can be reduced or adjusted, which can significantly decrease the cost of procuring and installing the wall. Such a buttress can also serve as a short-term or long-term remedy in case of observed excessive deformation in a constructed wall. Traditional methods of analyzing the stability of a sheetpile wall do not account for the passive pressure effects of this small buttress. In this paper, a parametric study was performed for a variety of soil types and wall and buttress geometries by implementing traditional limit equilibrium methods in commercial software and design charts were created illustrating the effect of a buttressing berm on sheetpile deflection. Selected cases were then modeled in a finite element software to compare the results with the traditional limit equilibrium method. The resulting design charts can be used as a tool by geotechnical engineers looking to reduce the design geometry of a sheetpile wall or address observed excessive deformation in a constructed sheetpile wall."

Jan 1, 2016

By B. H. Cody

THE Morenci concentrator of the Phelps Dodge Corporation is a part of the project that was completed after five years had been spent in development of the open-pit mine, equipment and process testing in the former No. 6 concentrator, and construction of the reduction works. The reduction works, of which the concentrator is a part, is about a mile and a quarter southeast of Morenci, on the highway that leads up Morenci Canyon. The concentrator has a nominal capacity for the treatment of 45,000 tons of ore daily. The original plant, which has a rated capacity of 25,000 tons of ore daily and is now in its fifth year of operation, began production on Jan. 30, 1942. The addition, to increase the capacity 8o pct, began operation on Dec. 1, 1943. The equipment of the original plant has been described in MINING AND METALLURGY (May 1942). No changes in equipment were recommended for installation in the concentrator extension and only one major change, forced by a shortage of maintenance labor, has since been made. Metallurgical treatment has been altered in some ways as experiments progressed and as labor and economic conditions dictated. No complete description of the operation of the concentrator has been published, so this paper will record such operating and equipment changes as have been necessary, the operating methods that have been developed and the metallurgical results that have been obtained. Descriptions are given in the order of the processes with a flowsheet and an equipment list of each process for convenient reference. The questions asked by engineer visitors have guided the detail in which operation and maintenance are described. During the period Jan. 30, 1942 to Sept. 1, 1946, ore milled has totaled 43,595,253 dry tons, from which 1,643,885 tons of concentrate have been made, containing an indicated production of 783,164,809 lb of copper. The principal results obtained in operation of the concentrator during the years 1942 to 1945, inclusive, and to Sept. 1, 1946 are shown in Table 1. A graphic general flowsheet of the plant is given in Fig I. ORE The Morenci ore that is now being treated is a medium-hard monzonite porphyry in which the principal sulphide minerals are chalcocite and pyrite. A minor amount of covellite is present. Chalcopyrite and bornite have been observed under the microscope. The chalcocite occurs mainly as a coating on pyrite. Some of the coatings are so thick that the pyrite presents the appearance of being an inclusion in chalcocite. Some are less than one micron thick and are barely visible at 500 diameters magnification of a polished cross section. Some chalcocite is present as a dissemination in the gangue. Various oxidized copper minerals are present as a result of oxidation in place but they represent a minor proportion of the copper minerals of the ore body. Oxidation

Jan 1, 1947

By F. O. Garrabrant

POWER is furnished to the Climax Molybdenum Co. by the Public Service Co. of Colorado over two 100,000-volt line to a bank of three 3333-kva. Transformers 100/13.8 kv. These transformers arc so designed that with the addition of forced cooling their capacity may be increased 33 1/3 per cent at 12,000-ft. altitude. Under present conditions, these transformers are operated at 140 per cent of their rating with forced air cooling. On the secondary side of this bank there is a 15-kv., 1200- amp. Westinghouse oil circuit breaker provided with automatic reclosing relays allowing three successive short circuits before lockout. From this breaker 13.8-kv. power is transmitted to a central switching station (Fig. I) where all circuits radiate to the different operations of the property. Each circuit is protected by a 23-kv., 600-amp. oil circuit breaker with by-pass and isolating disconnecting switches. MINE Power at 13.8 kv. enters the mine through a ventilation raise over a three-conductor, size 1-0 varnished cambric insulated, leaded, steel-wire armored cable, insulated from 17 kv. This power is received and distributed through one switching station and four underground substations, shown in Fig. 2. No. 370 substation (Fig. 3) contains three 250-kva, 13,800/440-volt, single-phase Pyranol transformers protected on the high-voltage side by a 600-amp. manually operated oil circuit breaker. This station serves 20 slusher-hoist motors, a 45-kw., three-phase, 220-volt electric furnace, a Massco hot miller for tempering and milling Timken detachable bits and a hoist with 175-hp. motor on the No. 3 development shaft. Also at this shaft are three pumps. One two-stage pump, 1200 gal, per min., 100-ft. head, 3600 r.p.m., driven by a 150-hp. motor, is at the bottom of the shaft. Two deep-well pumps, 500 gal. per min., 500-ft. head, are at the collar of the shaft. Under normal operation the one pumping 1200 gal. per min. is controlled by a float switch to take care of the water. If this pump fails, or if its capacity is insufficient, the first 500-gal. pump is automatically started, and if still more capacity is needed the second 500-gal. pump automatically starts. Time delay relays are provided on the deep-well pumps, so that if power fails the column will have time to drain before the pumps restart. On failure of any of these, an alarm is sounded in the train dispatcher's office, so that mechanics can be sent to ascertain the cause of failure and make repairs. The cable from No. 370 substation is run to the switching station, in which are four15-kv., 600-amp. direct-current solenoid-operated oil circuit breakers. The current for operating and tripping is furnished by a 125-volt storage battery. All breakers are operated from the switchboard in the compressor room. One circuit from this switching station goes to the

Jan 1, 1944

By N. H. Fisher

GEOGRAPHICAL DISTRIBUTION THE most important known deposits in Australia of what are commonly referred to as the beach-sand minerals are along the most easterly part of the Australian coast, between Southport, 17 miles north of the Queensland-New South Wales border, and Ballina, 50 miles south of the border, and most of the production has come from this area (Fig 1). Smaller deposits are known to occur farther to the south at intervals for several hundred miles, and beaches have been worked at Yamba, 96 miles south and Woolgoolga, 150 miles south of the border and at Swansea, 60 miles north of Sydney. Important deposits have been shown to exist on north Stradbroke Island and others have been located farther to the north, as far as Tin Can Bay at the south end of Frazer Island. (Fig 2). Table I gives the list of operators with the location of their deposits and workings, their approximate maximum monthly capacity expressed in tons of heavy mineral produced by their concentrating plants before separation into individual mineral concentrates, their methods of mining and concentrating the minerals, and the products obtained. This table applies as of June 1948, but all the operators are remodeling or improving their plants and plant practices will be modified accordingly. In addition, active boring campaigns have been carried out by Alluvial Gold, Ltd., of Sydney, in the Cudgen-Cudgera area, and by Zinc Corporation. Ltd., of Melbourne, on Stradbroke Island, and in other places. PHYSIOGRAPHY Submergence of the order of 100 to 200 ft at the close of the Pleistocene period left the pre-existing hills as promontories and the valleys as deep inlets. Wave action has built bay-bars and sandspits in a northerly direction from the headlands, and lakes and marshes have formed between these bars and the initial post-submergence coast line. These lakes and swamps have been or are being filled in by river-borne sediments or wind-blown sand. Beasley1 has presented evidence that suggests a recent emergence of the coast line of about 10 ft. The evidence for this belief is the presence of a black sand seam 1/3 mile inland and 17 ft above sea level (Fig 3), and although it is not entirely certain that this seam is of beach formation and owes nothing to windblown concentration, other evidence supports the suggestion of recent emergence. Such emergence would undoubtedly aid the formation of bars and sandspits and the easterly progress of the beach front, leaving a series of parallel dunes (Fig 4). This belt of north-south coastal dunes ranges up to 1/2 mile in width and as many as 15 lines of dune have been counted (Fig 3). The foredune is generally the highest, up to 25 ft high, and there is in many places a flat platform in front of the dune known as the berm. The dunes behind the foredune are lower and less well defined

Jan 1, 1947

By C. P. Newman, T. Gray, G. Beale, T. Cluff

Predictive geochemical and limnologic modeling of pit lakes is an important aspect of modern permitting. One of the key assumptions of many predictive pit-lake models is that highwalls and pit backfill become unreactive once submerged by the filling pit lake. Existing pit lakes provide useful data to test this assumption. The Sleeper pit lake (northwestern Nevada) is near hydrologic equilibrium and contains good-quality water that generally meets regulatory requirements for pit lakes. Despite the overall stable geochemical composition, seasonal trends in the hypolimnion indicate the generation of solutes and fluctuations in pH. This work applies mass balance, analysis of subaqueous pyrite oxidation, and trends in solute concentrations to evaluate the potential causes of long-term solute generation in the hypolimnion of the Sleeper pit lake. Results will aid in refining the effect that inundation of highwalls and waste rock has on ultimate pit lake chemistry and long-term closure planning. In the case of the Sleeper pit lake, ongoing solute loading does not substantially change the closure plan at this time as influent groundwater contains abundant alkalinity. INTRODUCTION As mine pit lakes become more widespread, regulatory agencies are increasingly requiring advanced predictive modeling of pit-lake geochemistry during permitting. The end goal of all those involved in these permitting actions (i.e., the mining company, the regulatory agencies, and the public) is to have pit lakes become geochemically and hydrologically stable for the long term, which may then allow mining permits to be relinquished. The modeling endeavors therefore are important to inform long-term closure of pit lakes. The conceptual models and numerical methods used to predict pit-lake water quality are complex and require substantial amounts of data and a variety of assumptions (Castendyk et al., 2015). One of the common modeling assumptions is that solute generation via acid mine drainage is eliminated (or at least decreased) by the filling of the pit lake and inundation of highwalls and any in-pit waste rock. However, few studies have reported on observations that support or refute this assumption. The water-quality record of the Sleeper pit lake is an ideal dataset with which to test these assumptions. The former Sleeper mine is located in northwestern Nevada (Fig. 1), and has been inactive (in terms of ore extraction) since 1996. The mine extracted Au and Ag from the famous high-grade Sleeper vein and associated widespread stockwork mineralization (Nash et al., 1995). Water quality in the pit lake has been monitored since the cessation of dewatering in 1996. The current pit lake actually occupies three separate sub-pits, the Sleeper, Wood, and Office pits (Fig. 1). The Sleeper and Wood pits undergo regular water-column monitoring using a conductivity-temperature-depth (CTD) probe, and sampling for dissolved inorganic constituents at several depths (~10, ~130, and ~330 ft. below lake surface). Groundwater quality is also monitored using a series of wells (Fig. 1).

Jan 1, 2019

By Wilfrid R. Foster

The industrial importance of the sillimanite group of minerals depends upon the beneficial properties exhibited by porcelains and refractories in which substantial amounts of these minerals are utilized as raw materials. The presence of "sillimanite" (now known to be the closely related mineral, mullite) in conventional porcelains and refractories has been known since before the turn of the century. As early as 1906 German investigators observed a direct correlation between the "sillimanite" content of a porcelain and its electrical and mechanical properties. Acting upon this discovery, U. S. Bureau of Standards investigators during World War I developed improved pyrometer tubes and spark plug porcelains, by increasing substantially the amount of "sillimanite" at the expense of quartz. The success of these experiments, first reported in 1918, at once aroused a lively interest in the location of natural sources of sillimanite and its related minerals. As a result of diligent search, the large andalusite deposits in California were discovered in 1919. Within the next several years large deposits of kyanite and sillimanite in India and kyanite in the United States had been reported. The search for these minerals in quantity, by both private and governmental agencies, has continued unabated during the ensuing forty years. A number of domestic deposits of kyanite have been worked during the past 30 years. Of these, only the Virginian and South Carolinian deposits are currently in substantial production. The only commercial andalusite production in the United States has been in California and Nevada. A dumortierite deposit in Nevada, known since 1917 but ignored for want of a commercial use, began production in 1925. A topaz occurrence in South Carolina has yielded a small production since about 1940. Domestic production of andalusite, dumortierite and topaz had virtually ceased by 1950. There has never been any commercial development of domestic sillimanite deposits. Kyanite has been mined almost continuously in India since about 1922. The remoteness of the large deposits of Indian sillimanite had caused it to be neglected until 1949. In that year sillimanite production began, and it has continued to the present. South African andalusite has been marketed from the 1930's until the present time. From 1944 to 1955 Kenya, British East Africa, was an important producer of kyanite. Commercial deposits of kyanite in Ontario, Canada should soon be in production. The first production of kyanite in Europe began in Austria in 1953. No commercial deposits of mullite are known anywhere in the world. It is of interest to note that synthetic mullite, which as "sillimanite" first elicited interest in the A12O3-rich aluminosilicate refractories, has enjoyed a marked increase in production in recent years. This is a result of the gradual depletion of known deposits of high-quality kyanite and related minerals."' Composition The sillimanite group includes kyanite, andalusite, and sillimanite, all having the same chemical composition (A12O3•SiO2) but different physical properties. Two other minerals, dumortierite and topaz, are also included in

Jan 1, 1960

By S. D. Warriner

ONE of the most, if not the most, difficult of the problems associated with the mining and distri-bution of anthracite is to get clearly into the minds of the consumers and of the editorial and reporto-rial writers in the public press the relationship between the cost of production, the price at the mine and the price that the consumer pays for his coal, delivered in the cellar. One sees in his morning paper a statement to the .effect that out of the $12 or $13 per ton that his dealer charges him, the miner gets only $2 or even less, and he asks himself or the paper asks for him, "Who gets the difference?" Not infrequently the amount that the miner gets is given for run-of-mine bituminous coal while the price quoted is that of prepared anthra-cite, egg, stove, or chestnut, as the case may be, delivered to the bin of the consumer; and the difference stands out with all the more prominence; and it sticks in his memory with a persistency which no manner of explanation is able to efface. A recently published official report for the year. 1921 of one of the State Bureaus stated the total production of anthracite in net tons of 2000 lb., and included not only the output recovered by washeries from old culm banks and by dredges from creek and river channels, but also the colliery consumption;, i.e., the coal used by the companies in the operation of the collieries. This resulted in a total production of 90,509,075 net tons for which the labor cost was stated at $283,961,300, or $3.13 per net ton, equivalent to $3.50 per gross ton. This figure of $3.50 was immediately seized upon as representing the labor cost per gross ton of fresh-mined commercial coal. It is almost a hopeless task to try to correct such errors of statement in the first place, and, in the second place, to correct the misapprehen-sions that arise from them. As a matter of fact, the report was misleading in regard to the labor cost of the entire production shipped, as the coal used for colliery consumption should be deducted. Correctly stated, the report should have shown, calculated in terms of gross tons of 2240 lb., the legal weight in Pennsylvania, as follows: Gross production 80,607,359 tons Less colliery fuel 8,757,137 tons Commercial production 71,850,222 tons Total labor bill $283,961,300.00 Labor cost per ton...: 8 3.952 The foregoing figures" of tonnage are obtained from the State Department of Mines. Even these include the quantity recovered from calm, banks by washeries, and consequently the $3.95 arrived at as the average labor cost is somewhat less than that actually involved in the mining of fresh-mined coal. Unfortunately, the figures of washery production and the cost of producing it are not available.

Jan 7, 1922

By Fred H. Poettmann, A. L. Benham

correlation of waters and identification of mixtures of two and three waters is important in secondary recovery evaluation studies and studies involving underground movement of waters. The importance of such studies suggests the need for a method to correlate a great number of water analyses to accurately identify and improve the understanding of mixtures of waters. Geochemical analyses of waters furnish a quantitative measure of each of the ions present in a water. Table L shows a portion of the data on many oilfield water analysis report forms in use today. Under "parts per million" are listed the quantitative amounts of each ion in milligrams per liter. Following the work of Stabler' and Palmer", some laboratories have reported additional data to indicate the reaction capacity, or reacting value of each ion. Table I lists data under two subheads for the "reacting value". Under the first subhead, the reacting value is reported in gram equivalents per million, hereinafter called equivalents per million. This value is determined by multiplying milligrams per liter of the ion by a factor equal to the reciprocal of the equivalent weight of the ion. Under the second subhead (per cent) is listed the percentage of each ion based on the total equivalents per million. Sodium is not usually an analyzed value in routine analyses, but is determined by the difference between total negative equivalent per million values and the sum of all of positive ion values for which analyses were made. This difference is reported as sodium, hereafter labelled "sodium". On this basis, the subtotals of the equivalent per million and per cent reacting values of the positive and negative ions are equal. CORRELATION OF WATER ANALYSES DATA Previously reported methods for correlating oilfield waters have used part of the data in Table 1, and can be classified into two categories. Correlation by Chemical Character Correlation methods using the chemical character of waters, as represented by the per cent reacting values of the ionic constituents, have been used by Palmer2 and Tickell. These methods have the limitation of neglecting the ionic concentration of the waters. This limitation can be serious in areas where waters have the same chemical character, but differ in concentration. Correlation by Concentration of the Ionic Constituents The Reistle' and the Stiff methods use the ionic concentration in parts per million and equivalents per million. These methods are effective in differentiating waters of similar chemical character and varying concentration, as in the highly saline waters of the Mid-Continent area. These methods have the disadvantage that they correlate waters by comparing many graphs of individual water analyses to determine the similarity to, or the variance from "type" waters. PROPOSED METHOD OF CORRELATING WATER ANALYSES DATA Correlation graphs presented in Fig. 1 represent water analyses data as points to facilitate the correlation of waters. The graph on the left of Fig. 1 is a plot of per cent reacting value of magnesium vs calcium. The right