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Beverley S. Randolph, Frostburg, Md. (communication to the Secretary): It has always appeared to me that each coalseam has its peculiar character or habit, due to the conditions of its formation, by virtue of which it is either persistent or variable in its section over extended areas. One seam will hold its thickness, chemical character, number and position of partings, beds of slate, etc., over a large area, while another will be found to vary in some of these characteristics every few feet. As the position of the outcrop is an accident of the topography, depending on folding and erosion which have taken place since the bed was formed, no one opening in the outcrop, or, in fact, anywhere else, can supply reliable evidence as to the general character of the seam. To determine the character of an untried seam, a number of sections are necessary from widely separated points, in order to show such variations as may exist. Should material variations be developed, more numerous sections are necessary, in

Jan 1, 1901

By Henry M. Howe

THE death of Sir Lowthian Bell removes almost the last of the group of heroic leaders who made their age and ours the Age of Steel-a group which his luster and the luster of his peers, Bessemer, Siemens and Tunner, made one of the most brilliant, perhaps even the most brilliant, in the annals of industry. Born on February 15th, 1816, in Newcastle-on-Tyne, he was the eldest son of Thomas Bell, an alderman of Newcastle, and of his wife Catherine Lowthian of Newbiggin, near Carlisle, in Cumberland, and was christened Isaac Lowthian, the name of his maternal grandfather. His father was born about 1788, and came in 1808 from Newbiggin to Newcastle, where he entered the office of the Cumbrian firm of Losh and Company, manufacturers of alkali and wrought-iron, at the Walker Iron and Alkali Works, at Walker, Dear Newcastle. Young Isaac Lowthian Bell was educated at Dr. Bruce's school in Newcastle, whence he went to Germany and Denmark, learning there German and Danish. On his return he studied at Edinburgh University, and then at the Sorboune in Paris. As a student he already showed the remarkable character and application which later gained him such distinction. After studying the manufacture of alkali near Marseilles, he returned, and when about twenty years old (A.D. 1836) was employed in his father's firm, then called Losh, Wilson, and Bell. After spending a year in their Newcastle offices, and after visiting the important iron-making districts of France, Germany, Poland, and Austria, he took charge of their rolling-mills at Walker, and profited so well by his opportunities, and especially by the teachings of his father and of Mr. John Vaughan (later one of the founders of the great firm of Bolckow, Vaughan & Co.), that he was made manager of the Walker works in or about

Sep 1, 1905

By Nicholas J. Grant, Rolf Nordheim, Bill C. Gissen

The system Cb-Re was examined in detail utilizing pure metals, careful melting techniques, and heat treatments. Metallographic and X-my methods were utilized for phase identification. In addition to soli-dus determinations, the composition limits and mode of formation of the intermetallic compounds 0 and X were determined. The binary constitution diagram Cb-Re has been the subject of a number of investigations.'-7 Two intermetallic phases have been described and tentative diagrams have been proposed. It was the aim of this investigation to obtain a complete and more accurate diagram through the use of purer alloys and improved techniques. The diagram worked out by Knapton7 was published after our experiments were concluded. EXPERIMENTAL METHODS The starting materials were Re powder of 99.95 pct purity, supplied by the Chase Brass and Copper Co., and Cb rondelles of 99.6 pct purity, supplied by the Electro Metallurgical Co. Typical analyses are given in Table I. The alloys were prepared from compacted Re powder and Cb rondelles; for critical compositions master alloys were used. The melts varied from 5 to 50 g and were melted nonconsumably in an Heraeus arc furnace, under ti-tanium-gettered laboratory grade argon. It was necessary to invert the buttons and remelt, followed by crushing and further remeltings, the number of which depended on the nature of the fracture of the alloy. The weight of each alloy was carefully checked after each melting operation. It was found that significant weight losses occurred only when loose Re powder was used. These losses amounted to a maximum of 3 wt pct in such instances, but did not occur after the charge was once molten, regardless of the composition. Typical loss of weight after 4 remelts, including crushing, amounted to 0.2 wt pct. Oxygen pickup during successive melting cycles did not appear to be significant, since the weight gain values never exceeded 0.01 wt pct. Control of the composition of the alloys was not too difficult; however, the problem of homogenization was a far greater one. Optimum alloying conditions were finally achieved by a combination of master alloy production and repeated crushing and remelting cycles. The compositions of the alloys utilized to determine the constitution diagram are listed in Table 11. The weight balance method was used ultimately as the sole analytical method, since repeated checks indicated that it was accurate to 0.2 pct, if the weight losses were small. The results were double-checked by a metallographic study of the homogeneity of the samples, since this was presumed to be the more serious factor. Heat treatments at 1200°C or lower were accomplished by sealing the specimens in Vycor tubes, heating in a Globar furnace, followed by air cooling. This was considered to be an adequate quench because of the very low temperatures involved. For temperatures of 1200" to 1900°, alloys were homogenized and heat treated in a tungsten-filament resistance vacuum furnace. From 1900" to 2750" a tantalum-tube resistance furnace was utilized, operating under a vacuum of 105 mm Hg. The temperature measurements up to 1200" were made by means of platinum-platinum 10 pct Rh thermocouples, with an accuracy of 5°C. For temperature measurements greater than 1200°C, optical pyrometers were used. They were calibrated against a standardized instru-

Jan 1, 1962

By Orson Cutler Shepard

RECENT research work in the flotation concentration of minerals has been concerned mainly with flotation reagents and the mechanism by which collecting reagents are held to the surface of certain mineral particles. The fundamental question of why air bubbles cling to properly prepared mineral surfaces is usually dismissed with the statement that the surface is air-avid or water-repellent. In my article in MINING AND METALLURGY for June, 1932, I showed theoretically that surface tension relationships govern bubble attachment in Rotation. This explanation has not been accepted by many; probably be¬cause the difficulty of measuring solid-solution surface tensions prevented obtaining experimental evidence. If the surface tensions involved in the system [solution, mineral, air bubble] are responsible for attachment of the air bubble to the mineral, any change in the surface tensions should cause a corresponding change in the attachment of the air bubble. An example in which the mineral-solution surface tension could be varied was found with mercury. When mercury is in contact with solutions of certain electrolytes the surface tension at the contact of mercury and solution can be altered by electrical polarization of the mercury. The surface tension rises to a maximum as the mercury is given an increasing negative potential and then decreases with further increase in potential. The theoretical relation between the surface tensions and the attachment of an air bubble is given by the following equation:

Jan 1, 1936

By AIME AIME

BECAUSE of its unequalled skill, your country in- creased its production until, in 1913, it produced 40 per cent. of the world's consumption of coal, iron ore, and cast iron; 45 per cent. of the steel; and 60 per cent. of the copper. Your industrial effort, in a general manner and your mining effort in particular, has been tremendous. In the general economy of the world, you occupy an incomparable position of power and action. All your curves of production, mineralogical, metallurgical, financial, commercial, have taken terrific trigonometrical tangents and you will easily under- stand the sorry position of your competitors. I must, therefore, pay a tribute of homage arid admiration to you all who, each in his sphere, have been the artisans of such prosperity and my earnest wish is that you continue in this way.

Jan 1, 1920

By E. W. Murbach

The oxidation of uranium carbide by oxygen at various pressures, and by air, has been investigated at temperatures up to 600°C. Arc-melted and cast uranium carbide displays oxidation behavior that appears to be dependent on its treatment after melting. ccReactive" uranium carbide, which results from exposure of freshly melted material to laboratory air, ignites in oxygen at about 275° to 350°C depending on pressure, and in air at about 350°C. The product of the initial oxidation appears to be UO2+.. Continued heating produces U30,. The initial oxidation rate is poportional, roughly, to the square root of the oxygen pressure. The activation energy determined for the initial oxidation reaction is 7.1 kcal, from 400° to 600°C in 10 mm of oxygen. AN investigation of low-decontamination methods for reprocessing uranium carbide reactor fuels is being carried out at Atomics International. One of the reprocessing cycles includes oxidation of the fuel at relatively low temperature (300° to 500°C) and reaction of the product oxide with carbon at elevated temperatures (above 1500°C) to produce uranium carbide. A number of small-scale experiments have been carried out in which uranium carbide was oxidized in air and oxygen at various pressures to measure reaction rates and to obtain other information concerning the oxidation process. In initial experiments with arc-melted UC, ignition was observed at about 275°C in oxygen or 350°C in air. However, similar experiments by Neymark' at 426°C in air showed no ignition. It was soon determined that the difference in oxidation behavior was due to the treatment of the UC following arc-melting. Neymark's experiments were carried out with freshly melted UC while those reported in the

Jan 1, 1963

By Frank C. Roberts

Neglecting the hydrogen and hydrocarbons, 1 will assume the following analysis as a fair average composition, by weight, of the waste gases escaping from a coke-burning blast-furnace: CO2............17.25 CO,............25.25 N,.............57.50 100.00 The 25.25 CO requires for combustion 62.6 of air, or 11.398 of oxygen, necessitating the passage of 48.202 nitrogen into the products of combustion. The latter, therefore, have the following composition: CO2,,............ 56.9 N,............105.7 Tile total specific heats of these quantities of the respective gases are 12.31 and 25.79, giving for the combined product, 38.1 approximately. That is, the product of combustion will absorb 38.1

Jan 1, 1889

By G. E. Anderson

THE Asientos-Tepezala district is in the north of the State of Aguascalientes, about 30 miles north of the city of Aguascalientes, the capital. The district is reached by a standard-gage railway on the west side through the junction, Rincon de Romos, which is on the main line of the National Railways of Mexico from El Paso to Aguascalientes. The terminal is Tepezala. On the east side a narrow gage connects Asientos with the junction, San Gil, on the main line, Asientos to San Luis Potosi. Tepezala on the west side of the Asientos hills is about 2 miles from Asientos on the east side. The area between the two mining centers is separated by a low subued divide which forms the central part of the hills. The district is in a group of isolated hills surrounded by characteristic bolson plains. The hills rise above the surrounding plains about 6500 to 10,000 ft.; the highest peak is Alta Mira. The two mining districts lie somewhat upon the flanks of the hills at about the same elevation, 7200 ft. The district is an old one and has been known since early in the 17th Century when the mines were worked by the Franciscan Monks. Their mining operations which were limited to the oxidized ore above the water level, consisted chiefly in extracting silver from the Santa Francisca vein. To judge from the old workings and the remains of surface plant, the operations were on a large scale. By modern mining operations since about 1900, the various mines of the district have been one of the principal sources of ore for the smelter located at Aguascalientes.

Jan 2, 1926

By G. P. Conard, R. C. Hall, J. F. Libsch

The 50 pct Fe-50 pct Co alloy undergoes a transformation from disorder to an ordered structure of the CsCl type reportedly in the vicinity of 732OC. During this process, the coercive force goes through a maximum, apparently as a result of strains associated with the coherent nucleation and growth reaction. This magnetic alloy also shows a marked increase in the ratio of residual to saturation induction, which is associated with annealing to a high degree of order with the continuous application of a magnetic field. The increase in ratio can be explained on the basis of a decrease in 90' domain boundaries and, perhaps, by an increase in anisotropy resulting from lattice distortion. THE 50 pct Fe-50 pct Co alloy undergoes a disorder-order transformation which has been reported to occur in the vicinity of 732°C1,2 The ordered structure is the CsCl type.' This magnetic alloy also shows a marked increase in the ratio of residual to saturation induction as a result of heat treatment in a magnetic field, sometimes called a response to magnetic anneal.'-' The purpose of this investigation was to study the course of the ordering reaction, the nature of the response to .heat treatment in a magnetic field, and the relation, if any, between ordering and the response. Procedure The method of approach in this investigation was to produce an initial structure as completely disordered as possible and then gradually to order the alloy by isothermal anneals at various temperatures under different conditions of the applied magnetic field. Magnetic, magnetostriction, and X-ray analyses were of primary importance in determining the property and structural changes resulting from the isothermal anneals. Rings of the 50 pct Fe-50 pct Co alloy were prepared from the elemental powders by a powder metallurgy technique, further details of which may be found in ref. 7. The initial structure was produced by annealing the specimens for ½ hr at 1000°C, cooling to and holding for ½ hr at 900°C (in the a range above the ordering temperature), and water quenching. Isothermal anneals were performed at 600°, 675°, 720°, and 740°C. For example, rings were heated to 600°C, held for a predetermined period of time, and cooled by natural cooling at a rate slightly slower than an air cool (average of 20" to 25°C per min). The tests (magnetic, etc.) were made after each heat treatment. All high temperature treatments were performed in a purified hydrogen atmosphere. The treatments at the various temperatures were carried out under one or more conditions of an applied field including 1—no field, 2—field of 20 oersteds applied on cooling only, and 3—field of 20 oersteds applied continuously during heating, holding, and cooling. Magnetic measurements were made using the standard Rowland ring technique8 with a maximum field strength of 100 oersteds. The magnetization curve, induction at 100 oersteds (B.), residual induction (Bt), and coercive force (Hc) were determined. All magnetic analysis data were based on an average of the results from three rings. A strain gage technique9 as used for the measurement of magnetostriction. The X-ray determination of the relative amount of ordered phase present was made on the ring specimen used for magnetic measurement. This was done by the back-reflection method using a rotating specimen (because of the large grain size) with unfiltered CoKa radiation and a 7 hr exposure time. Intensity measurements of the ordered line (300) were made by comparing visually the films so obtained with standard films prepared by exposing for different lengths of time a specimen given a long time anneal (high degree of order). Results In all instances the saturation induction (induction at 100 oersteds) was found to increase slightly with annealing time. This effect was small and appears to be the increase in saturation induction to be expected on ordering.10-13 The residual induction behavior was markedly influenced by the field condition during annealing, Figs. 1, 2. For the condition of no applied field, the ratio of residual to saturation induction remained essentially constant for short annealing times but showed a significant increase at longer times. With increasing annealing temperature, less time was required to produce this increase in the ratio. In the case of the 600°C anneals, the increase did not occur until approximately 20 hr, Fig. I, while on annealing at 740°C the increase was immediate, Fig. 2. Slight decreases in the ratio may be observed at 100 hr for specimens treated at 720°C and at 1 hr for those treated at 740°C. Specimens annealed in a field of 20 oersteds showed a residual to saturation induction ratio consistently higher than that for the specimens annealed without the field. The first anneal with the field (¼ hr) caused an abrupt increase in the ratio at all temperatures; thereafter, the increase in the ratio was generally similar for specimens annealed

Jan 1, 1956

By Charles Mitke

(San Francisco Meeting, September, 1915) INTRODUCTION THE Copper Queen mine is composed of seven divisions which are operated through the following shafts: Division Shaft Depth, Air Current No. Feet Uncle Sam 600 Downcast (shut clown April, 1914, to 1 curtail production) Southwest 600 Downcast 2 Czar 400 Downcast 3 Holbrook 600 Downcast 4 Spray S00 Downcast (shut clown April, 1914, to curtail production) 5 Gardner 1,000 Downcast 6 1 Dallas 1,400 Downcast 1 Lowell 1,600 Upcast 7 Sacramento 1,700 Downcast " I I The workings of the different shafts are connected by motor-haulage drifts on the even numbered levels. The general location of the orebodies and workings is illustrated by the vertical projection of orebodies of the district as shown in Fig. 1. The Uncle Sam, Southwest, and Czar workings have many connections to the surface through raises and extensive cracks that were caused by moving ground. These divisions are cool and are ventilated entirely by natural means. The Holbrook is ventilated partly by natural and partly by artificial ventilation, while the Spray, Gardner, Lowell; and Sacramento are ventilated entirely by mechanical means. In some of the divisions mentioned above a considerable quantity of air exhausts through shafts of adjoining properties.

Jan 9, 1915

By Leon Dominian

(Wilkes-Barre Meeting, June, 1911.) I. INTRODUCTION. THE lack of Aryan roots for the names of metals commonly known among the Aryan settlers of Asia Minor, as well as the later colonizers of Europe, indicates that these races were generally ignorant of the use of metals until they came into contact with Semitic peoples. Practically all mining-terms in current use among the earliest Greeks resemble very strongly their distinctly Semitic equivalents, which can be traced all the way in a broad belt beginning in Lower Mesopotamia, and extending westwardly to the Syrian shores of the Mediterranean. The Greek word " "metallon," for instance, used indiscriminately to designate mine or ore, probably came from the earlier Semitic equivalent, " matal." Again, the Greek words " chrysos " (gold) and " chalkos " (copper) seem to be descended from the Semitic forms " throats " and a chalak." It is a natural inference that primitive mining-methods were evolved by the dwellers in the mineralized areas of Asia Minor, from whom later Greek, Roman, and even North European miners obtained their first notions of the reduction of metallic ores, by virtue of a general westward migration of mining and metallurgy. Some traces of its passage through Turkish territory will be noted in this paper. While European Turkey can boast of one ancient gold-field, the Asiatic dominions of the Sultan may lay claim to at least two well-defined and widely-separated gold-producing districts. These three regions may be distinguished as the Thracian, the Pontic, and the Anatolian gold-fields. II. TURKEY IN EUROPE. . The Thracian Gold-Field. The most conspicuous topographic feature of the lowland between Constantinople and Salonica is the uplifted Archaean

Nov 1, 1911

By Edward A. Vitunac

The emerging importance of pneumatic conveying systems for bulk materials is largely due to the economic advantages such systems frequently exhibit when compared to alternate methods. Included in these advantages are relative simplicity, ease of installation, flexibility, ease of maintenance, cleanliness, safety and ease of automation to the degree desired. In general, installed costs favor a pneumatic system when compared to competing mechanical systems if the transporting distances are long or to and from many points. In addition, supplemental use such as cooling, heating, or drying may supply the incremental value required to justify its selection.1 The main disadvantages of a pneumatic system are the greater horsepower required per unit weight of material conveyed, material limits in abrasiveness, and density. In addition, material breakage tends to increase at changes of direction and at any obstruction such as that caused by the imperfect butting of two pipe sections in the conveying line. Classification of the material at the discharge may also cause problems in dusting and segregation. Materials weighing less than 4 lb per cu ft tend to cause feeding and collecting problems although they can be conveyed easily. Material between 4 lb and 20 lb per cu ft can be handled with fewer problems, but flow rate is hard to predict. The 20-60 lb per cu ft range offers best results, whereas materials of over 60 Ib per cu ft tend to drop out of the air stream and settle in the pipeline unless careful pro- visions are made to prevent this.2

Jan 5, 1968

By J. W. Goss, M. J. Coolbaugh

Most grouting has been done to stop water flaw in mines and for stabilizing foundations of various man-made structures, a survey of the U.S. literature reveals. Apparently Sun Manuel is one of the first mines in this country to use grout extensively underground for strengthening and stabilizing ground in drifts, shafts, and stations. A comparison with other procedures as well as details of the Sun Manuel program are covered. The employment of pressure grouting as a ground stabilizer at the San Manuel Copper Corp. mine reduces delays in both development and production, lowers costs, and makes possible safer working conditions. It specifically reduces delays in haulage operations and permits the maintenance of normal ventilation. In development work this grouting allows faster excavation by cementing together highly fractured or broken ground that otherwise would require extensive cribbing or spiling. In drift repair, it consolidates the loose or fractured rock over the timber or steel drift supports, thus decreasing the frequency of repair, lessening the hazards from falling rock, and curtailing delays due to blockage of drifts by muckpiles and repair operations. Pressure grouting is the process of pumping an accurately controlled mixture of cement and water into loose, fractured, or porous rock. The ratio of water to cement varies according to the nature of the rock encountered, from the thinnest mixture of 30 gal of water per sack of cement, used to fill very fine fissures, to very thick grout of 5 gal of water per sack of cement, used to fill large fissures or extensive areas of loose, broken rock. The pumping pressures at San Manuel vary from about 100 to 1000 psi, depending upon the compactness of the rock, while at other mines the pressures sometimes go as high as 5000 psi. In order to grout an area in which ground water is encountered, the pumping pressure of the grout must be increased by an amount equal to the pressure of the ground water. GROUTING IN DEVELOPMENT WORK The areas under development at San Manuel include many sections of loose, highly fractured rock. Prior to the advent of the grouting program, drifts, shafts, and stations in these areas could not be excavated without utilizing extensive support such as cribbing, spiling, and breast boarding. This slowed down the development work and increased the costs considerably over that required for excavating in more competent ground. Now, when a particularly bad area is contacted, the mining is temporarily stopped while the bad portion of drift or shaft is grouted. After the grouting is completed, the excavating is resumed with an approximate decrease of 50 pct in lost time and costs. Grouting in Drifts and Turnouts: The grouting procedures and patterns used in drifts and turnouts are very similar. The standard procedure for grouting such a turnout is shown in Fig. 1. After the ground is supported as well as possible with cribbing or lagging, the gaps between the back lagging and side lagging are plugged with empty cement sacks or additional timber, if necessary. The ground surrounding the proposed turnout is then grouted in two stages. In the first stage, the broken and caved rock is drilled and grouted to a depth of 5 to 10 ft, depending on the depth of the broken rock. This stage forms a grouted seal that allows higher pressures to be used at depths beyond 10 ft without developing excessive leaks at the face. After a grout hole is drilled, a 10-ft long pipe is wedged tightly into it with empty cement sacks. Grout is then pumped into the hole until the pressure reaches 200 to 300 psi, and the pipe is subsequently removed if it hasn't become cemented in. No more than one hole at a time is drilled and grouted because the grout has a tendency to go from one hole into another, plugging up the latter. Five or six holes are usually adequate for the first 10 ft of grouting. Experience has shown bentonite to be a useful admixture to the grout, particularly when it is indicated that the grout is being lost into large fissures or voids. Bentonite increases the plasticity of the grout enabling it to remain in place more easily until it has begun to set. In the second stage, the longer holes are drilled through the previously grouted rock, after which the

Jan 1, 1961

By Henri V. Reibell

Deep open pit optimization supposes very long and sedious calculations in order to assign the best shape of the pit and the best bottom level, which will give the biggest profit. Computers give the possibility to make these calculations. More over, they allow to multiply trial calculations so that it is possible to obtain more reliable results. The paper does not expose the computer program but only the way in which mining engineers can use it.

Jan 1, 1969

By Raymond E. Zimmerman

Abstract-The Swirl Cyclone developed in Japan, although based upon the same principles as the standard cyclones for heavy-media or water-only cyclones for cleaning coal, is novel in that it is invested with the apex on top, discharging rejects at the top, and clean coal at the bottom. It also uses a variable air vent tube for controlling the vortex. Tests conducted indicate that its efficiency is not only as sharp as standard cyclones but it also performs using a much coarser grain magnetite, promising higher magnetite recovery. The author describes the results of tests conducted in Japan on a very difficult coal.

Jan 2, 1978

By Karl J. Benecke

Though the first patent on a bucket wheel excavator (BWE) was granted in 1881 in the US, this technology was developed in Germany to the high standards of today. However, this development was only possible in parallel with development of shiftable belt conveyors and boom stackers. The combination of these three types of equipment in the open pits of the Federal Republic of Germany has made lignite the cheapest energy source for power generation, despite the low heating value of German lignite on the order of 8500 kJ/kg (3600 Btu/lb) only. Comparing the costs for fuels used in power generating on the basis of their heating value, the relation between lignite, high quality black coal, and imported crude oil is on the order of 1 to 1.5 to 2.5, hence highly in favour of lignite.

Jan 8, 1979

By E. C. Houston

The phosphate ores mined in middle Tennessee typically consist of granular rock phosphate particles disseminated in a clayey matrix. In the TVA plant near Columbia, Tenn., the phosphate ore is mined, made into a slurry with the addition of a small amount of sodium hydroxide as dispersant, and treated in a hydroseparator to remove minus 10 micron material. The hydro-separator underflow, comprising a rough concentrate, is transported by pipeline to the plant; the hydro-separator overflow, comprising a tailing, is flocculated by addition of calcium sulphate and is discharged to settling ponds at a rate of about 1400 gpm. Sedimentation in the ponds produces a clarified effluent, which may be recycled for use as process water or discharged to surface drainage. This method of tailing treatment is not entirely satisfactory since poor sedimentation characteristics of the tailing result in poor ultimate utilization of pond storage volume.

Jan 1, 1949

By E. C. Wright, P. W. Mumma

This paper describes some properties of steels in the composition range 0.10 to 0.30 per cent carbon and 1 to 7 per cent chromium. It is well known that some steels of this type develop high tensile strengths upon receiving comparatively simple heat treatments, such as ordinary air cooling from appropriate temperatures, with or without subsequent tempering treatments. These high strengths are frequently associated with unusual toughness as measured by the Izod or Charpy impact test. The chromium content, in addition to imparting the above properties, enhances the resistance to corrosion and improves the strength and oxidation resistance at elevated temperatures. Moreover, these steels have good forging and fabricating properties, as evidenced by the fact that several thousand tons of seamless tubing have been made during the past five years from steels containing 0.10 to 0.25 per cent carbon and 5 per cent chromium. Composition and Properties of Materials Tested Twelve 15-lb. induction furnace heats of the compositions shown in Table 1 were prepared, using low-carbon open-hearth steel base melts

Jan 1, 1933

By Wm. C. Madge, H. P. Allen

The chloridizing mill of the Standard Reduction Co. is located about 75 miles south of Salt Lake City on the Tintic branch of the Denver & Rio Grande Western R. R. and 12 miles from the Tintic Standard mine. The daily capacity is 200 toss of a siliceous, low-grade, silver-lead ore from this property. It has operated continuously since it was started in January, 1921. The process consists essentially of a chloridizing roast followed by a percolating leach with a nearly saturated solution of common salt, acidified with sulfuric acid, the precipitation of silver on sponge copper and of copper and lead on tin-plate cuttings. The precipitates are shipped to a smelter. Some of the general ideas involved are said to have been used by Augustin in England, in 1840. A number of textbooks treat of the subject, especially the chloridizing roast followed by a leach with sodium hyposulfite or amalgamation. The process was revived in this district by Theo. P. Holt, N. C. Christensen, the Bureau of Mines, and others. Nature of Ore Treated The average assay of the ore treated during 1924 is as follows: Gold, ounces per ton...................................... 0.025 Silver, ounces per ton..................................... 18.26 Copper, per cent......................................... 0.30 Lead, per cent........................................... 5.00 Silica, per cent........................................... 65.00 Iron, per cent............................................ 10.00 Lime, per cent........................................... 0.70 Sulfur, per cent.......................................... 3.00 Arsenic, per cent......................................... 0.70 The silver is finely disseminated and occurs as native, combined as a sulfide and, to a very small extent, as the chloride. The lead may be present as carbonate, sulfide, or sulfate.

Jan 1, 1926

By A. E. Lee

THE Blackwell Zinc Co., Inc., a subsidiary of the American Metal Co., Ltd., operates a horizontal retort zinc smelter at Blackwell, Okla. The plant has 14 furnace blocks of 800 retorts each, fired with natural gas on a 48 hr cycle. Over 13,000 tons of zinc-bearing material, chiefly sulphide flotation concentrates, are treated monthly to produce slab zinc and high lead-cadmium fume. In 1942 a program of rebuilding and modernizing the smelter was started. By 1947 furnace smelting capacity had been increased to a point where roasting and sintering facilities were inadequate, and it was necessary to purchase oxidized materials to supplement sinter production. The seven 210 ft Ropp roasters and three 42 in. x 44 ft Dwight-Lloyd machines then in use had been in service at least 20 years and were in need of major rebuilding. Thus it was entirely practical to consider all new equipment and a change of method rather than rebuilding and repairing obsolete units. A study of the problem indicated that roasting as such could be eliminated and roasting and sintering accomplished in one step by a modification of the Robson process,' which had been used since the early 1930's by the National Smelting Co., Ltd., at their plants at Avonmouth, England, and Swansea Vale, South Wales. Francis P. Sinn, General Manager, Zinc Smelting Operations, The American Metal Co., Ltd., who was familiar with the practice in England, suggested the use of one large machine for the entire operation from concentrate to sinter. One step sintering appeared to best meet Blackwell's plant requirements and indicated substantial savings in labor, gas, coal, and repair costs. Choice of Machine Size The sinter machine size was set at 12x168 ft for a rated capacity of 540 tons per day. This tonnage, produced on a five day week, would meet the seven day requirements of the 14 furnace blocks. The one large machine was quoted at a lower cost than two or more 6 ft wide machines of similar total capacity. Further, the larger machine could be housed in a smaller structure and only one set of equipment for charge preparation and delivery and for disposal of sinter cake was needed. One machine on a five day week made possible a concentration of the skilled operating personnel and required less men than a plant including two or more machines and related equipment circuits. Fewer units of equipment meant less maintenance, and the two down days weekly allowed ample time to repair and, if necessary, to make up lost production. Experience had indicated better sintering quality and rates with larger masses of material, not only on wider machines, but also in deeper beds. The ratio of windbox perimeter to area for the 12x168 ft machine is 0.179, compared to 0.353 for a 6x102 ft machine and 0.617 for a 42 in. x 44 ft machine. This meant less air leakage with resulting fan power savings and less spoilage of charge along the pallet sides. Performance Initial operation of the new sinter plant was made in November 1951 and regular production attained late in December. The average product sinter output during 1952 and the first half of 1953 has been 18.2 tons per hr. The average for one month has been as high as 22.4 tons per hr. Considerable experimenting with varied operating conditions accounts in part for the below capacity — 24 tons per hr — average output, and work to further improve production rate continues. A typical sinter analyses is 66.0 pct Zn, 0.3 pct Pb, 0.1 pct Cd, 0.3 pct S, 8.0 pct Fe, 2.0 pct SiO,, 0.8 pct CaO, and 0.2 pct MgO. Use of this material has made possible increases in furnace burden and improved furnace operation over the former practice using sinter made from Ropp roasted concentrates. Better lead and cadmium elimination in sintering has permitted the furnace production of slab zinc lower in lead and cadmium. Anticipated economies of operation have largely been gained. The sinter plant is operated by seven men per 8 hr shift — one head operator, three equipment operators and three sweepers — plus one oiler on day shift only. While it has been necessary at times to operate seven days a week to produce the required sinter tonnage, the five day work week usually has been adequate. Consumption of natural gas for sinter bed ignition is 200,000 to 300,000 cu ft per day. Green Ore Sintering Practice The 30 to 31 pct sulphur content of the —200 mesh zinc concentrates is the fuel used to sinter the charge, no coal addition being required. In the feed to the machine, sufficient concentrates are added to crushed return sinter fines containing 0.3 to 0.5 pct sulphur to produce a charge averaging 5.0 to 6.5 pct sulphur. Since the return sinter used in Blackwell's practice is varied from — 1/2 to — 1/8 in., the actual sintering mixture of fine sinter and concentrates is somewhat higher in sulphur. The coarser sinter particles are too large to resinter and merely aid porosity in the sinter bed. The ratio of concentrates to return sinter in the charge ranges from about 1:4 to 1:5.5. Variations are based on the appearance of pried up bed sections, bed exit gas temperature trends, windbox suctions, and return sinter size. Sufficient sulphur must be used to obtain fritting of the charge into a soft sinter cake and to aid in the elimination of lead and cadmium. Excessive feed sulphur will result in partial slagging of the cake impairing porosity and prolonging sintering time.

Jan 1, 1954