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By C. B. Manula, R. Venkataramani, T. V. Falkie

A linear programming model has been developed to solve a production planning problem for a multi-plant stone operation serving several customers. The factors considered are production capacities, inventory restrictions, demand requirements, and available stone reserves. The objective function to be optimized is the total annual net returns from the quarry operations. For the purposes of model testing, annual stone sales from each plant were divided monthwise, productwise, and assigned to a market. The concept of assigning demands to market sectors and not to any particular plant creates the self-correcting interplay between the plants and the markets. Along with the inventory restrictions, and ending inventory requirements, this interaction enables the model to minimize the external diseconomics. The model was tested against an operating prototype and a number of alternative problems were also considered. Finally, the basic linear programming model has been enlarged by the addition of another dimension and the problem has been derived in a decomposition format.

Jan 1, 1972

By Arthur F. Johnson

THE principal costs of making a pound of aluminum are for 9 kw-hr of electricity and for 1.9 Ib of the oxide (A1203) called alumina. A pound of alumina is made by digesting 2 or 3 lb of bauxite in hot caustic solutions, settling and filtering out the oxides of silicon, iron, and titanium and precipitating and calcining the A1203 as a pure, white powder. The cost of foreign bauxite delivered at an alumina plant in the U.S. may equal or exceed the cost of its chemical treatment to make alumina. Some of the principal factors affecting costs in the utilization of foreign bauxites are: transportation from mine to market, mineralogical and physical composition of the bauxite, thickness of the bauxite beds and their clay and forest overburden, amen- ability of the bauxite to beneficiation and to caustic extraction of the alumina, and lastly size of bauxite deposits and rate of mining employed.

Jan 6, 1954

By Harold Bloom, H. E. Hawkes

Streams and rivers are the principal channels into which the weathering products of rocks and their contained ores are funneled. The inorganic load of a stream system is a crude sample of all the earth materials in the drainage basin of the stream, and the heavy mineral assemblage of sediments is one of the oldest and most successful guides to certain kinds of bedrock ore in unexplored terrain. Early application of this method was made possible by the ease of separation and identification of heavy minerals in the field with the simplest equipment. Water analysis for traces of base metals became a practical possibility about eight years ago with the development of rapid chemical tests that could be carried out at the field site. Since that time the water method has been widely used in mineral reconnaissance, and a substantial number of discoveries have been reported. Here again, the techniques developed for this work were relatively simple and could be performed easily at the field site.

Nov 1, 1956

By K. V. S. Sastry, M. Cross

A summary of recent progress in understanding the main factors affecting the pelletization of fine particles is presented. A scheme for categorizing the mechanisms responsible for pellet growth is out- lined and laboratory methods to both identify the important growth mechanisms and quantify the effects of the major factors influencing the rate of pelletization are described. Next laboratory and mathematical simulations of the commercial process involved in pelletizing iron ore concentrates, are presented. The operational instabilities associated with this pelletization process are described and subjected to a control theory analysis. This analysis, in conjunction with the mathematical models, is used to both explain existing control problems and assess the effectiveness of possible circuit redesigns at producing a more stable and controllable operation. Although much of the discussion is in the context of iron ore pelletizing the approach developed to analyze the agglomeration behavior of these concentrates should apply to most other fine particle systems.

Jan 1, 1980

By R. F. Shurtz

This paper relates geostatistical doctrine to classical statistical and mathematical concepts. It is not a contribution to geostatistics; it is a contribution about geostatistics. It uses descriptive and pictorial arguments to assist interested readers in forming an intuitive grasp of the recondite and difficult mathematical issues involved. The method chosen is to cite authoritative literature references from the two fields and trace connections between them. Four broad conclusions appear: first, geostatistics comprises a relabelled subset of concepts detached from classical statistics and mathematics; second, this relabelling artificially isolates geostatistics from the classical fields and restricts its scope; third, the classical fields still have much to offer to the mining industry but will remain fallow unless artifcial communication barriers are removed; and fourth, the unique advantages asserted for geostatistics are neither unique nor necessarily real. These findings should not be regarded as iconoclastic but rather as Pointing the way to a wider and more productive scope for those who evaluate mineral deposits.

Jan 1, 1983

By C. P. Proctor, J. T. Crawford

TROUGH washers were among the earliest methods used for concentrating ores; they are referred to by Agricola about the middle of the sixteenth century as already being used while the hand- operated jigging sieve had only recently come into use. For cleaning coal the trough washer apparently was first used in France and Belgium about 1841. The early trough washers were intermittent and the refuse at least was shoveled out by hand. The invention of a continuous jig washer, by a Frenchman, Berard, in 1848, acted as a stimulus to coal washings in jigs and retarded progress in trough washing in the nineteenth century. Chapman and Mott' describe a number of trough washers developed in England and on the Continent from 1850 to 1900, such as the Bell and Ramsay, Ramsay, Bell and Green, Wunderlich, Scaife, McLellan, Lodge, Elliott, Blackett, and Murton. The later of these removed the refuse mechanically, either intermittently or continuously, but in ways that interfered with stratification.

Jan 1, 1943

By Dean W. Lynch

Arizona: The San Xavier Indian Reservation mine is located approximately 2 ½ miles northwest of Asarco's Mission Mine southwest of Tucson. The company plans to produce oxide and copper-bearing siliceous ore at this operation, construction of which is scheduled for completion in 12 months. Asarco also plans a copper leaching and iron precipitation plant at its mine. The plant will leach 4000 tpd of oxide ore using sulfuric acid produced at the company's acid plant now under construction at Hayden. Copper precipitates will go to the smelter at Hayden. A drilling program at the Sacaton copper prospect in Pinal County, has disclosed two porphyry copper deposits. One deposit, mineable by surface mining, has approximately 17 million tons of material averaging 0.80% Cu. A deeper deposit suitable for under- ground mining has approximately 12 million tons averaging 1.5% Cu. Although Asarco intends to place this property in production, no startup date has been determined.

Jan 1, 1970

By D. E. Ranta, M. W. Ganster, A. D. Ward

An ore zoning technique developed by Climax geologists during the past forty years has been used to contour drill-hole assay values on cross-sections for fourteen porphyry molybdenum deposits in western North America, including the Climax and Henderson orebodies. Detailed geological studies of these large, complex deposits are essential in determining ore controls, which are the factors most critical to the zoning of MoS2 and to conceptualizing geologic model s. The technique provides a detailed representation of the size, shape, grade distribution, and geologic continuity of the individual deposits. It demands a disciplined approach and adheres to rigorous rules while maintaining flexibility for adaptation to widely varying geologic conditions. By defining the limits that guide and control manual, computer-assisted conventional, and geostatistical reserve estimates, it exerts a strong geologic influence on the estimation process. A combination of ore zoning and geostatistical analysis of molybdenum deposits produces a significantly more accurate, precise, and defendable geologic reserve estimate than other methods.

Jan 1, 1984

By Peter N. Calder, G. H. Blackwell

The development of an open pit slope monitoring system, from equipment selection and justification to complete computer data storage and analysts, Is described. Methods of overcoming the Limitations of the electronic transit/distance measuring device and survey system selected are discussed. Results of the monitoring program showed that blasting was a major cause of Instability, and, after much experimentation, suitable blast patterns for mine production and pit wall integrity were designed. The experiences gained with double benching 'and wall cleanup procedures are discussed . The monitoring program also indicated that water was a contributor to instability, and methods of dealing with water In a rock with a low storage coefficient and transmissibility are described. Because of the danger of falling rock, photogrammetric methods are used for joint mapping and analysis, this work being contracted out. Joint strengths are measured on-site using a portable shear tester. Finally, a discussion of a complex failure is given.

Jan 1, 1983

Dr. Rausch, from your vantage point as educator, engineer, and corporate executive, what would you say is the most fundamental need that must be satisfied for the continuing welfare of our society, our industry and our profession?

Jan 3, 1977

By Henry H. Logan

THE purpose of mine-hoist safety controllers is to prevent hoists from being operated at speeds above those considered safe and expedient, to ensure proper acceleration and retardation near the ends of travel, and to prevent overwinding. In other words, they are to safeguard the normal speeds and movements of the hoisting drums and cages. The object of this paper is to explain the fundamental principles of hoist controllers, and to aid in their selection, installation, and maintenance. There have been many designs of mechanical hoist controllers, the basic elements of which as ultimately developed consisted of a flyball governor and another movable member supported by a frame and arranged to move in a vertical, horizontal, or rotary direction, depending upon the particular design, the whole mechanism arranged to be driven in synchronism with the hoist drum. In case of overspeeding, inadequate retardation, or overwinding, a pawl or catch controlled by the governor would engage teeth or dogs on the moving member, which through mechanical connections would cut off the power and set the brakes. Since in strictly mechanical hoist controllers valuable time was lost while the various mechanical movements took place, an important development was the invention of controllers partly mechanical and partly electrical. This type of controller has a governor and movable cams to be driven in synchronism with the hoist drum, the governor representing the hoist speeds and the cams profiled to form a pattern for retardation, representing the travel of the cages in the shaft. This type is also provided with an overspeed switch and usually with a pair of selective-limit switches, one for the top limit and the other for the bottom. In overspeeding or failure to retard in accordance with the profile of the cam, the overspeed switch will be opened, or if the top or bottom limit is overrun one of the limit switches will be opened, breaking the retaining circuit of a contactor or magnetic latch, causing the power to be cut off and the brakes applied. When the speed falls to the predetermined safe speed, the overspeed switch closes, and a hoist equipped with power-operated brakes controlled in emergency by a solenoid-operated valve can be restarted immediately by simply moving the master control lever to the "off" position and then applying power-unless brake interlocks are used, in which case the brake lever or levers must also be brought to the "applied" position, after which the hoist can be restarted in the usual way. When hand-reset weights instead of solenoid-operated mechanism are arranged to set the brakes for emergency, it is necessary to reset the weights manually, besides bringing the master control lever to the "off" position before the hoist can be restarted. Should a final limit be slightly overrun, and a stop made with a limit switch held open by a cam, it is necessary for the hoistman to use a "backing-out" switch which short-circuits the overrun limit switch, reestablishing the hoist-control circuit, and simultaneously opening the

Jan 1, 1940

By L. D. Jaffe

IN 1945-1946, the author measured temper brittle-ness in ingots of varying composition prepared by remelting SAE 3135 bar stock under vacuum. Since other investigators1-" have been referring to this hitherto unpublished work, belated publication seems worthwhile. Steel in 7 lb batches&apos; was induction melted with nickel additions in a magnesia crucible under a pressure of 10 microns of Hg or less and cast into a 2 in. diam chill mold. The compositions of ingots A, B, and C prepared in this way are shown in Table I; the vacuum melting lowered carbon, manganese, sulphur, and nitrogen content. Ingot B was homogenized 24 hr at 1095°C; ingots A and C were not homogenized. Blanks, M in. square, were cut from each ingot, austenitized 1 hr at 870°C, water quenched, tempered 1 hr at 595"C, and water quenched. Half the blanks from each ingot were given an embrittling treatment of 50 hr at 455°C and water quenched. A few specimens, embrittled and unembrittled, were finally heated 1 hr at 580°C and water quenched. Charpy specimens machined from the blanks were V-notched on the side closest to ingot mid-radius and paired as to radial position in the ingot, one of each pair having received the embrittling treatment and the other not. Specimens were broken in a standard impact machine at various temperatures, the same temperature, in general, being used for both bars of a pair. Fig. 1 shows the results. It is evident that temper brittleness, as measured by the increase in temperature of transition from brittle to tough failure introduced by the 455°C treatment, was decreased but not eliminated by reducing carbon to 0.006 pct, manganese to < 0.004 pct, and nitrogen to 0.001 pct. In another study," temper brittleness was apparently eliminated by -reducing carbon to 0.003 pct, with 0.80 pct Mn and 0.0015 pct N (ingot D)." • Analyses given for carbon below 0.02 pct were made by Jensen method; for nitrogen, by Kjehldahl method. Values quoted else-where&apos;-&apos;." for carbon content of ingot C and for nitrogen content of ingot D were obtained by conventional combustlon and by vacuum fusion methods, respectively, and are considered much less accurate. The 580°C final treatment removed previous em-brittlement (ingot B particularly), indicating that this was "reversible embrittlement."" Metallographic examination of ingot C with 1 pct nital etch revealed only ferrite; ingot A had carbide spheroids uniformly distributed in a ferrite matrix. With ethereal picric acid plus zephiran chloride etchant.&apos; no differences between embrittled and un- embrittled specimens of ingots A, B, or C were noted. Fracture of C and D below the transition temperature was predominantly transcrystalline, with traces of intercrystalline, in both embrittled and unembrittled conditions. The corresponding fracture of unembrittled B was mixed transcrystalline and intercrystalline; of embrittled B, predominantly intercrystalline with some transcrystalline areas. . . . Acknowledgment The efforts of J. C. Leschen, formerly of National Research CO~P., in preparing the ingots are acknowledged with thanks.

Jan 1, 1956

By H. J. Hager, P. H. Haskell

For the past 20 years Alabama has probably led all other coal-producing districts in the proportion of coal prepared by wet washing. All rail mines, with one exception, and a high percentage of truck mines now wash all coal under 3 or 4 in. Most of this coal has been cleaned by jigs, though within the past few years one or more wet-table plants have been installed for cleaning some of the coal from practically all the major seams. These tables supplement either single or multicell jigs usually for one of the following purposes: (I) to wash the finer sizes of raw coal, (2) to rewash the finer sizes of washed coal after they leave the jigs, (3) to re-treat jig middling products. The main purpose of this paper is to present cleaning data and operating factors involved in the preparation by tables of the finer sizes of raw coal. The specific information and data pertain to the Mary Lee seam of coal as extracted at three of the mines of the Alabama By-Products Corporation, which produces coal for both byproduct coke-oven consumption and the general steam market. The Mary Lee seam has long been recognized by coal-preparation authorities as presenting a very difficult ash-reduction problem. Washing characteristics and physical structure of the seam vary widely. Its thickness ranges from 32 in. with no rock parting to 100 in. with from one to three partings. Over the entire field the bed contains bands of bone and rash. These are distributed throughout the seam cross section. A portion of this bone or bony coal occurs as definite bands I in. or more thick; however, some bone is interlocked as layers within the coal and consequently breaks into flakes when the coal is broken. This material is relatively high in ash and its specific gravity varies widely. It has been observed that in most cases in the washing of Mary Lee coal from 10 to 18 per cent of the raw coal is within +0.10 specific gravity of the attempted point of separation. Therefore with the average wet jig, only with extreme difficulty can a low-ash product be obtained without loss of much marketable coal in the refuse. Without entering into the controversial question as to whether modern washing equipment can efficiently handle simultaneously all sizes up to a given maximum, suffice it to state that the comparatively old-type jigs in service at the Alabama By-Products Corporation&apos;s mines could not clean the he sizes satisfactorily. Therefore, in 1931 that company installed six diagonal-deck tables at the Barney mine, which were installed as an addition to a complete plant cleaning all coal below minus 3 in. The primary washed product is shipped as steam coal and another portion may be prepared as washed nut for domestic trade. These tables are washing 9 tons per hour of raw coal sized to ¼ in. to o. The washed table coal is mixed with the 3 to ¼-in. jig product to constitute a 3 in. to o steam coal.

Jan 1, 1942

By H. J. Hager, P. H. Haskell

For the past 20 years Alabama has probably led all other coal-producing districts in the proportion of coal prepared by wet washing. All rail mines, with one exception, and a high percentage of truck mines now wash all coal under 3 or 4 in. Most of this coal has been cleaned by jigs, though within the past few years one or more wet-table plants have been installed for cleaning some of the coal from practically all the major seams. These tables supplement either single or multicell jigs usually for one of the following purposes: (I) to wash the finer sizes of raw coal, (2) to rewash the finer sizes of washed coal after they leave the jigs, (3) to re-treat jig middling products. The main purpose of this paper is to present cleaning data and operating factors involved in the preparation by tables of the finer sizes of raw coal. The specific information and data pertain to the Mary Lee seam of coal as extracted at three of the mines of the Alabama By-Products Corporation, which produces coal for both byproduct coke-oven consumption and the general steam market. The Mary Lee seam has long been recognized by coal-preparation authorities as presenting a very difficult ash-reduction problem. Washing characteristics and physical structure of the seam vary widely. Its thickness ranges from 32 in. with no rock parting to 100 in. with from one to three partings. Over the entire field the bed contains bands of bone and rash. These are distributed throughout the seam cross section. A portion of this bone or bony coal occurs as definite bands I in. or more thick; however, some bone is interlocked as layers within the coal and consequently breaks into flakes when the coal is broken. This material is relatively high in ash and its specific gravity varies widely. It has been observed that in most cases in the washing of Mary Lee coal from 10 to 18 per cent of the raw coal is within +0.10 specific gravity of the attempted point of separation. Therefore with the average wet jig, only with extreme difficulty can a low-ash product be obtained without loss of much marketable coal in the refuse. Without entering into the controversial question as to whether modern washing equipment can efficiently handle simultaneously all sizes up to a given maximum, suffice it to state that the comparatively old-type jigs in service at the Alabama By-Products Corporation&apos;s mines could not clean the he sizes satisfactorily. Therefore, in 1931 that company installed six diagonal-deck tables at the Barney mine, which were installed as an addition to a complete plant cleaning all coal below minus 3 in. The primary washed product is shipped as steam coal and another portion may be prepared as washed nut for domestic trade. These tables are washing 9 tons per hour of raw coal sized to ¼ in. to o. The washed table coal is mixed with the 3 to ¼-in. jig product to constitute a 3 in. to o steam coal.

Jan 1, 1942

By W. C. Knoblaugh

Rotary kilns are adaptable to many fuels, but this paper deals principally with the use of powdered coal. The observations and conclusions presented are based on rotary kilns used in the manufacture of basic refractories for steel furnaces. Solid fuels are pulverized for two principal reasons: (I) ease of control, and (2) rapidity of combustion. By the use of pulverized coal, the operator can increase or decrease the fuel supply by small increments and with very little time lag between the change of control and the resultant change in the flame. This is extremely important in this type of furnace, as the temperature range for maturing the clinker is very narrow. By fine grinding, it is possible to increase the rate of burning of the coal, thereby liberating great quantities of heat within a small space in a short time. Under given conditions, temperature is a measure of the heat liberated per unit of time. To attain the temperatures of 2950° to 3150ºF., which are encountered in the manufacture of various steel-furnace refractories, it is estimated that it is necessary to liberate 125,000 to 175,000 B.t.u. per cubic foot of the burning zone per hour. The manufacture of dolomitic refractories is a ceramic process and, as is true in most such processes, the material must be subjected to a certain critical temperature for a certain length of time in order to complete the reaction. If the time of reaction is reduced, the temperature must be increased; in other words, if the rate at which the material flows through the hot zone of the kiln is increased, the temperature of the hot zone must be increased to compensate for this shorter exposure to the flame. Firing with Pulverized Coal Fig. I is a line drawing of a rotary kiln with the attendant equipment indicated in their relative positions. The kiln proper consists of a sloping, refractory-lined steel tube rotating slowly about its longitudinal axis (0.75 r.p.m.), and is countercurrent in operation in that the raw feed is introduced at the stack end and gradually approaches the zone of highest temperature at the lower end, where the fuel is introduced. Before 1930, the usual method of firing a rotary kiln was to store the coal in the pulverized form and then feed from this storage to the kiln as needed. This proved unsatisfactory and dangerous because of dust explosions and the inability to obtain an even feed to the kiln caused by arching of the coal in storage and flooding of the coal feeders. In later years the practice has been to store the raw coal and grind it as it is needed. Such a layout is indicated in Fig. I. The coal is stored in bins, charged from the top and drawn from the bottom into recording scales and then into some type of unit pulverizer. There are a number of good pulverizers .on the market, each incorporating features that the manufacturer considers outstanding.

Jan 1, 1948

By H. I. Aaronson, C. Wells

THIS is a preliminary report on a detailed metal- lographic investigation of the bainite reaction in a plain carbon hypoeutectoid steel. In view of the current interest in the bainite reaction and of the controversies which still exist about even the basic phenomenology of this mode of transformation, it was considered desirable to present a summary of the more important results obtained to date prior to the completion of this investigation. The steel used contained 0.29 pct C, 0.76 pct Mn, 0.25 pct Si, 0.005 pct P, and 0.007 pct S. The alloy was homogenized for 48 hr at 1250°C. Individual specimens were then austenitized for 30 min at 1300°C and isothermally transformed for various times at temperatures above and within the bainite region, ranging from 775" through 500°C at intervals of 25°C. These specimens were carefully examined on a research metallograph; a few specimens of especial interest were also studied electron microscopically.&apos; Bainite may be defined as the aggregate structure resulting from the precipitation of carbides from supersaturated ferrite which formed by nucleation and diffusion-limited growth. Two groups of experimental observations made during the present study provide support for this mechanism of the bainite reaction: 1—at 625°C, partial conversion of pro-eutectoid ferrite to bainite began after more than 99 pct of the austenite had decomposed and 2—the precipitation of bainitic carbides occurred only within the ferrite phase. The proeutectoid ferrite and the bainite reactions in the steel studied were found to be morphologically continuous. Following the classification scheme of Mehl and Dube,2 three principal morphologies of proeutectoid ferrite were identified in the temperature region (625" to 550°C) in which the two reactions blend—grain boundary allotriomorphs," * Grain boundary allotriomorphs are crystals which nucleate at and grow preferentially along matrix grain boundaries. Widmanstaetten side plates, and intragranular plates. Carbide precipitation converted these morphologies into their bainitic equivalents with increasing isothermal reaction time. Conversion took place more rapidly as the reaction temperature was reduced. The replacement of supersaturated ferrite by bainite, however, did not occur uniformly even within individual austenite grains and was noticeably incomplete at higher transformation temperatures, e.g., 625° to 575°C. The ferritic and bainitic forms of a given morphology thus were frequently found to coexist, often inextricably intermixed, in a given specimen. At 625" and 600°C, the conversion of ferrite to bainite was clearly seen to have occurred entirely by the nucleation of carbides at ferrite/ferrite boundaries, as shown in Fig. 1. The nature of the loci at which carbides formed at lower temperatures was often difficult to ascertain with a light microscope, although the alignment in rows of many of the carbides, illustrated in Fig. 2, suggested that nucleation had again occurred at grain boundaries or subboundaries within the ferrite. Application of the electron microscope confirmed this deduction for bainitic structures formed at these temperatures, as shown in Fig. 3. The dense black appearance of bainite formed at very low reaction temperatures in higher carbon steels3 indicates, however, that isolated lattice defects within ferrite crystals prob-

Jan 1, 1956

By Jon J. Connor, Richard J. Ebens

The trace element geochemistry of the cherty clay-rich carbonate residuum cover in much of southern Missouri and adjacent parts of Arkansas was investigated during the earl 1970&apos;s as part of a statewide geochemical survey of Missouri.1-8 The study was environmentally oriented and focused on the character of the "natural" geochemical environment of the state. A great deal of effort was made to collect samples that would represent the geochemical variability of the near-surface environment in an unbiased way. In general, this was accomplished by use of randomization procedures in both collection and analysis. Most of the sampling sites were selected randomly to prevent personal bias from entering into sample selection. In the laboratory, samples were analyzed in a randomized sequence to circumvent the effects of any systematic bias that may have entered into the laboratory procedures. Several residuum samples contained 1000 parts per million (ppm) or more of lead, zinc, and barium. All these samples were collected in Washington County in or near the barite district. The frequency distributions shown in [Fig. 1] indicate just how unusual the residuum in Washington County is with regard to these three metals and copper. The shaded parts of the histograms represent the samples from Washington County. The histograms are based on analyses of 141 samples from 38 localities scattered over the subcrop of the Cambrian and Ordovician formations of southern Missouri and northern Arkansas. Eight localities were selected randomly from over each of four bedrock units-Bonneterre dolomite; Potosi and Erninence dolomites; Roubidoux formation; and the Jefferson City, Powell, and Cotter dolomites. Six localities were selected randomly from over the Gasconade dolomite. The geochemical "background" for these elements in residuum is defined by samples from outside of Washington County (unshaded part of the histogram); lognormal distribution curves have been fitted to the corresponding parts of each histogram. Clearly, these lognormal models (curves) describe well the background distributions of zinc, barium, and copper. The background histogram of lead concentrations is notably skewed to the right compared to the model. Lognormal distribution curves were also fitted to the data on lead, copper, and zinc in the residuum of Washington County. No model was fitted for barium in Washington County because of the distinct bimodality in the data. All samples containing more than 1000 ppm were collected in the barite mining district and obviously reflect the elevated levels of barium in the district. The frequency distributions for Washington County are clearly anomalous with respect to the background data, those for lead, zinc, and barium being exceptionally so. The models for anomalous lead, zinc, and copper seem to fit reasonably well, although the scatters in the lead and zinc anomalies are considerably larger than the scatters in the corresponding background data. The copper anomaly, although probably real, is greatly subdued compared to that of lead and zinc. Also, it has a scatter similar to that of the copper background, rather than being enhanced like the scatters of the lead and zinc anomalies. On the basis of techniques described by Miesch,4 the chemical abundance of copper, lead, and zinc in residuum in

Jan 1, 1980

By W. C. Knoblaugh

ROTARY kilns are adaptable to many fuels, but this paper deals principally with the use of powdered coal. The observations and conclusions presented are based on rotary kilns used in the manufacture of basic refractories for steel furnaces. Solid fuels are pulverized for two principal reasons: (I) ease of control, and (2) rapidity of combustion. By the use of pulverized coal, the operator can increase or decrease the fuel supply by small increments and with very little time lag between the change of control and the resultant change in the flame. This is extremely important in this type of furnace, as the temperature range for maturing the clinker is very narrow. By fine grinding, it is possible to increase the rate of burning of the coal, thereby liberating great quantities of heat within a small space in a short time. Under given conditions, temperature is a measure of the heat liberated per unit of time. To attain the temperatures of 2950° to 3150°F., which are encountered in the manufacture of various steel-furnace refractories, it is estimated that it is necessary to liberate 125,000 to I75,000 B.t.u. per cubic foot of the burning zone per hour. The manufacture of dolomitic refractories is a ceramic process and, as is true in most such processes, the material must be subjected to a certain critical temperature for a certain length of time in order to complete the reaction. If the time of reaction is reduced, the temperature must be increased; in other words, if the rate at which the material flows through the hot zone of the kiln is increased, the temperature of the hot zone must be increased to compensate for this shorter exposure to the flame. FIRING WITH PULVERIZED COAL Fig. I is a line drawing of a rotary kiln with the attendant equipment indicated in their relative positions. The kiln proper consists of a sloping, refractory-lined steel tube rotating slowly about its longitudinal axis (0.75 r.p.m.), and is countercurrent in operation in that the raw feed is introduced at the stack end and gradually approaches the zone of highest temperature at the lower end, where the fuel is introduced. Before 1930, the usual method of firing a rotary kiln was to store the coal in the pulverized form and then feed from this storage to the kiln as needed. This proved unsatisfactory and dangerous because of dust explosions and the inability to obtain an even feed to the kiln caused by arching of the coal in storage and flooding of the coal feeders. In later years the practice has been to store the raw coal and grind it as it is needed. Such a layout is indicated in Fig. I. The coal is stored in bins, charged from the top and drawn from the bottom into recording scales and then into some type of unit pulverizer. There are a number of good pulverizers on the market, each incorporating features that the manufacturer considers outstanding.

Jan 1, 1946

By W. C. Knoblaugh

Rotary kilns are adaptable to many fuels, but this paper deals principally with the use of powdered coal. The observations and conclusions presented are based on rotary kilns used in the manufacture of basic refractories for steel furnaces. Solid fuels are pulverized for two principal reasons: (I) ease of control, and (2) rapidity of combustion. By the use of pulverized coal, the operator can increase or decrease the fuel supply by small increments and with very little time lag between the change of control and the resultant change in the flame. This is extremely important in this type of furnace, as the temperature range for maturing the clinker is very narrow. By fine grinding, it is possible to increase the rate of burning of the coal, thereby liberating great quantities of heat within a small space in a short time. Under given conditions, temperature is a measure of the heat liberated per unit of time. To attain the temperatures of 2950° to 3150ºF., which are encountered in the manufacture of various steel-furnace refractories, it is estimated that it is necessary to liberate 125,000 to 175,000 B.t.u. per cubic foot of the burning zone per hour. The manufacture of dolomitic refractories is a ceramic process and, as is true in most such processes, the material must be subjected to a certain critical temperature for a certain length of time in order to complete the reaction. If the time of reaction is reduced, the temperature must be increased; in other words, if the rate at which the material flows through the hot zone of the kiln is increased, the temperature of the hot zone must be increased to compensate for this shorter exposure to the flame. Firing with Pulverized Coal Fig. I is a line drawing of a rotary kiln with the attendant equipment indicated in their relative positions. The kiln proper consists of a sloping, refractory-lined steel tube rotating slowly about its longitudinal axis (0.75 r.p.m.), and is countercurrent in operation in that the raw feed is introduced at the stack end and gradually approaches the zone of highest temperature at the lower end, where the fuel is introduced. Before 1930, the usual method of firing a rotary kiln was to store the coal in the pulverized form and then feed from this storage to the kiln as needed. This proved unsatisfactory and dangerous because of dust explosions and the inability to obtain an even feed to the kiln caused by arching of the coal in storage and flooding of the coal feeders. In later years the practice has been to store the raw coal and grind it as it is needed. Such a layout is indicated in Fig. I. The coal is stored in bins, charged from the top and drawn from the bottom into recording scales and then into some type of unit pulverizer. There are a number of good pulverizers .on the market, each incorporating features that the manufacturer considers outstanding.

Jan 1, 1948

By E. H. Crabtree, Neil S. Parker

THE custom milling plant of the Eagle-Picher Mining and Smelting Co. is at Sahuarita, Arizona, approximately 20 miles south of Tucson. It is connected by a 2-mile railroad spur to the main line of the Southern Pacific Railroad between Tucson and Nogales. The capacity of the plant is 500 tons per day. The mill is conventional except perhaps in the rather wide variety of ores treated. These vary from straight copper ores, from which only one concentrate is produced, and lead-zinc or copper-zinc ores from which two concentrates are produced, to complex gold-silver-copper-lead-zinc ores, which require the simultaneous production of three or four concentrates. Custom ores are received in lots varying in size from a few truckloads to several carloads. Similar ores, after crushing and sampling, may be commingled, or individual lots may be milled individually. The ores are first dumped either into one of two 200-ton railroad bins, or into one of the two 75-ton truck bins. From these bins the lot of ore is individually crushed in a 24 by 16-in. jaw crusher to 3-in. size, and then to 1/2-in. size in a 3-ft Symons shorthead crusher. After crushing to 1/2-in., the ore is conveyed to the automatic sampling plant. This consists of three Vezin samplers in series, each successively cutting out 10 pct, to pct and 5 pct of its feed. After No. 1 sampler, the material passes through a mixing barrel and is then crushed to 1/4-in. in a set of rolls before passing to No. 2 sampler. Between No. 2 and No. 3 samplers, the ore is again mixed in a mixing barrel. The final sample, consisting of one pound per ton of ore, is taken to the laboratory, where it is further crushed to 1/8-in. in a coffeemill, and then riffled down to the size required for grinding in a pulverizer. The capacity of the crushing and sampling plant is up to too tons per hour, depending upon the character of the ore. After sampling, the ore is stored in five ore-storage bins, consisting of four 200-ton bins and one 175-ton bin. Each of these bins is discharged by means of belt feeders and conveyor to the ball mills, so that ores may be composited for milling in any required ratio, or milled separately by themselves. Fine grinding is done in one 6 by 4-ft Allis-Chalmers ball mill and one 8-ft by 36-in. Hardinge mill. These two mills both discharge to one 54-in. Akins Simplex Highweir classifier. The classifier sands can be returned to either or both mills, so that flexibility in grinding for different ores is easily obtained. Flotation equipment consists of four 66-in. Fagergren machines for copper-lead roughing and six similar machines for zinc roughing. Lead-copper bulk concentrate is cleaned and recleaned in eight No. 18-S Denver machines, and zinc concentrate is cleaned in three 66-in. Fagergrens. Copper-lead differential separation

Jan 1, 1947