Search Documents

By Bhudeo N. Sinha, Paul Dean Proctor

Heavy metal contents in stream waters, sediments, and selected aquatic algae were determined for the upper Meramec River basin, Missouri, of 3905 km2 (1508 sq mi) area and site of the proposed and controversial Meramec dam and reservoir. Six distinct drainage areas within the basin are: (1) a lightly industrialized and farm area, (2) a primitive forested- ranch area, (3) an area which includes the Northern New Lead Belt, (4) the westward drainage of an old and extensive barite mining area, (5) the main trunk of the Meramec River, and (6) short north-divide tributaries of the Meramec River. Lower Paleozoic sediments rest nonconformably upon eroded and hilly Precambrian igneous basement. This locally rises as major inliers and also makes up the St. Francois Mountains to the east. Major lead, and lesser zinc, copper, cobalt, and nickel mineralization occurs in the buried Cambrian Bonneterre dolomite. Residual barite is present along the eastern fringe of the basin and is derived from the Potosi formation. Some 62 water samples, active stream sediments (AS), bank samples (BS), and 11 selected aquatic algae samples were collected in the various drainage areas and analyzed for copper, lead, zinc, cadmium, nickel, arsenic, and mercury contents. Heavy metals in water have a maximum range of 1-120 ppb (parts per billion), and this for zinc. The AS and BS sediments have orders of magnitude greater acid leachable metal contents in the -80 mesh fraction, with lead hawing the greatest range of 11-1028 ppm. Heavy metal content of stream-algae approaches that of the sediments in which they grow. Most anomalous heavy metal accumulations occur near mining and milling activities and in the streams draining such areas. Significant distortions of the primary dispersion pattern militate against geochemical exploration for hidden ore zones, but do demonstrate the mining, milling, and smelting activities of man, and sporadic mineral occurrences in younger horizons.

Jan 1, 1980

By W. H. Corbould

The township of Cloncurry is situated in the northwestern part of Queensland, about latitude S. 20" 42' 53" and longitude E. 140" 30' 25". Townsville is the port through which all the trade passes and is about 470 miles east-northeast of Cloncurry township. The Cloncurry copper district covers an area about 200 miles long by 40 miles wide. Unfortunately the mines are very scattered over this area, there being no large groups of mines. Mount Elliott mine is situated about 70 miles south from the township of Cloncurry, and Mount Oxide mine about 150 miles north of Cloncurry. The district has been known for the last 35 years, and in the early stages a large amount of capital was expended with little to show for it; but within the last four years the railway has been extended from Townsville to Mount Elliott and about 40 miles north of Cloncurry, which has made possible the treatment of the ores of the mines on this extension. The Mount Elliott company within the last few years has smelted and converted on its mine 169,688 tons of ore for 20,554 tons of blister copper, containing 20,377 tons of copper, 35,642 oz. of gold and 28,841 oz. of silver, at a cost of about 233 per ton of blister copper, f.o.b. Townsville. This cost included all development, wages, salaries, general management, office, agency, legal expenses, freight, etc. The gold and silver contents would practically pay for freight to Europe, refining, realization charges, etc. The costs of supplies are as follows: Firewood, 15s., coke, 75s., mine timber, 54s., and coal, 60s. per long ton; explosives: dynamite and gelignite, 50s., and blasting gelatine, 65s. 6d. per case, Mount Elliott Mine The pay shoot of the orebody outcrops at the surface for about 200 ft., and lies between two slate walls about 40 ft. apart., having a general

Jan 1, 1915

By Percival Roberts

I DESIRE to present this evening a few notes from the standpoint of a manufacturer upon the subject of tests of iron for structural purposes; not that I have any new facts to which to call your attention, but rather a statement of a few of the objectionable features of the present modes of operation, showing the absolute necessity of important and very radical changes. At the present time, as most of you are well aware, proposals for materials are accompanied by certain specifications as to quality and tests to which the manufactured article must conform. Evolved in many cases from the inner consciousness of some one more expert in theory than in practice, we find a few of the specifications to be good, many bad, and some indifferent. Appended are specimens of specifications of the most prominent engineers and railroad companies, which may be taken as representing the best and most intelligently framed specifications which we now have, arid yet in some pints scarcely satisfactory, owing to misunderstanding of tests as now made. To show the very varied opinions held as to proper quality of iron for bridge-work, I give the tests of two lots of fron made by inspectors with the results, which were referred to the maker with the following commensts : Lot No. 1. Teat of Iron for Eyebars. Reduced area. Elongation. Size. Per cent. Per cent. Breaking load. 4" x 1 3/8 26 25 68,500 Bent 120° around 1 1/2" bar with fracture. Reduced area. Elongation. Size. Percent. Percent. Breaking load. 6" x 1 7/15 24 24 48,820 Bent 110° around 2" bar. This iron should bend 180° without fracture, and is therefore rejected. Lot No. 2. Test of Iron for Brirlye. All measurements taken upon a parallel section of 10". Iron for tension members. Breaking load Elongation. size. per aquare inch. Percent. 1 5/8 49,600 20 14 61,600 27

Jan 1, 1882

By Robert Franke

(Butte Meeting, August, 1915.) IN view of the prominence which the conical mill has attained in the fine-crushing field within the few years since its introduction, the following comparison with its more mature forerunner, the Chilean mill, based on extensive tests, is submitted in the interest of the milling profession. Soon after designing the concentrating plant of the Miami Copper Co. in 1909, the Hardinge conical mill made its appearance in the milling forum. Its possibility as a suitable crushing device for the plant was well recognized, but in view of the lack of commercial demonstration, at the time, as to capacity, efficiency for desired product, and the still more uncertain factors of cost of maintenance and power consumption, it was deemed that the immediate adoption of this machine throughout the plant would be a hazardous undertaking. For these reasons it was decided to equip the majority of the imme diately required units of the plant with Chilean mills, the fine-crushing proficiencies of which were better known, and one section with Hardinge mills, to serve as a test unit for the guidance of future installations and replacements. Thereby, after 1.5 years' operation with both types of mills, a thorough test as to metallurgical efficiency and cost economy has been obtained. The conical mill used in these tests is the 8-ft. Hardinge pebble mill, having a cylinder 22 in. in length. The cylindrical portion of this mill is lined with cast-iron liner plates, and the conical extensions with silex bricks bound together by cement. Each liner plate carries a projecting lifter, the function of which is to increase the height of drop of the lifted material. Danish No. 5 pebbles, obtained from the coast deposits of Denmark, are used for the grinding charge. The Chilean mill used is a fast-running, 3-roller, 6-ft. Saturn mill, with screens of 0.037-in. opening. The feed to these mills is the oversize of Callow screens having 0.029-in. openings, which follow rolls crushing to 0.5 in.

Jan 7, 1913

By C. A. Heberlein

The present war in Europe seems to have stimulated the demand for quicksilver. In July last, the price ranged around $35 per flask of 75 Ib., while to-day it seems to fluctuate between $47.50 and $50. The use of quicksilver in the form of fulminate, containing 70 per cent., and as corrosive sublimate, containing 73.8 per cent., must be extensive and a further market advance may be expected. With such an advance, some revival in quicksilver mining may be looked for, particularly in California, where many abandoned properties exist. What can be done with low-grade quicksilver ores under favorable conditions, particularly with low mining cost, may be seen from the following description of the Oceanic mine, near San Luis Obispo, Cal. Its operations date back to 1876, when rich sand ores were discovered. In three years 7,400 flasks of quicksilver were produced, but since then the deposit has been worked only sporadically . Until two years ago various attempts had been made by the Oceanic Quicksilver Co. to mine the more refractory "mud rock," which was encountered lying beyond the " sand rock." Mud and sand rock in varying quantities were treated together, with poor results, but operations were successful with sand rock of fair percentage. About two years ago, when all the sand rock had been exhausted, Murray Innes purchased the Oceanic mine. He found large quantities of low-grade mud rock, varying from 6 to 8 Ib. of quicksilver per ton, which at the prevailing low price of quicksilver demanded very cheap mining costs, in conjunction with high but cheap metallurgical extraction. At that time the extraction of quicksilver from mud rock ranged between 40 and 50 per cent., as far as I am able to learn. Mining The low cost of mining was the principal consideration, since, with mercury at 50c. per pound, the total gross value of the ore did not exceed

Jan 1, 1915

By George H. Rupp

THE Sunrise iron mine of The Colorado Fuel and Iron Corporation is in Platte County, Wyoming, about 110 miles north of Cheyenne. It is served by the company-owned Colorado and Wyoming Railway, which connects with the Chicago, Burlington and Quincy and the Colorado and Southern railways at Guernsey, Wyoming, a distance of 6 miles from the mine. The original Sunrise claims were staked mainly by United States Army officers and soldiers stationed at old Fort Laramie, because of small showings of copper in the Carboniferous limestones that form a horizontal capping overlying the Huronian iron ore. A small copper smelter was operated for about three years at Fairbanks, now an abandoned town on the North Platte River. As many of the copper shafts and test pits near the present Sunrise mine were deepened, they penetrated the capping and encountered iron ore or gray sericitic schist, and were abandoned. Evidence of an early interest in iron ore was found when the Chicago claim was stripped for mining. A narrow, winding crevice was discovered, which led downward 20 ft. through the capping to the underlying soft red hematite, in which a room 10 by 14 by 7 ft. had been excavated. Stone hammers and wedges and a few rusted pieces of iron were found on the floor, and the evidence is conclusive that the excavation was used up to rather recent times by Sioux Indian medicine men, to provide war paint. GEOLOGY AND ORE The Sunrise district is in the Hartville uplift, an anticlinal area in which pre-Cambrian rocks similar to those of the Black Hills and the Laramie Mountains are brought to the surface. The lowest part of the pre-Cambrian series exposed is a great thickness of pink, crystalline, Huronian dolomites with thin unstratified beds of schist and quartzite. The upper part of the series consists mostly of gray or green schists but includes a few lenticular beds of jasper and a banded iron formation con-

Jan 1, 1939

By R. L. Templin

In treating a cast metal by any working process such as rolling, drawing or forging, variations in the conditions present in the remelting, casting, chilling and preheating of the initial ingot will certainly have some effect on the physical properties of the final product. In considering the effects of cold working on the physical properties of metals, however, it will be assumed that these factors, being controlled according to best practice, will remain constant and therefore may be neglected. Starting with such a uniform metal ingot, the next operation in the production of wrought metal involves hot working by rolling, forging or extrusion so as to produce a slab, billet or bloom. This working must be sufficient to produce substantial grain refinement and minimize any defects occurring in the initial ingot. Even such a hot-worked product is not suitable as a starting material for determining the effects of cold working because it contains internal strains analogous to those resulting from cold working. It is necessary, therefore, to supplement the hot working with a considerable amount of cold working, followed by a proper annealing in order to obtain a material suitable for use in defining the effects of cold working on the physical properties of a metal. That is, if the effects of cold working are to be evaluated accurately it is necessary to have other effects at least constant if not absent. Such material, unfortunately, cannot always be obtained in commercial practice but the requirements just indicated certainly represent the desired goal; this goal has been attained in the treatment of some metals, one being aluminum. Factors That Must Be Considered Simultaneously with Cold Working So many other factors affect the physical properties of a metal that it is necessary to consider these factors simultaneously with cold working; their effects vary in magnitude with the amount of cold work present. The most important of these factors are: (I) chemical composition, (2) structure, (3) temperature of working, (4) character of working, (5) rate of working, and (6) methods used for measuring results. For

Jan 1, 1930

By Ellsworth Daggett

The ore smelted in the Winnamuck furnace during the year 1872 consisted, for the most part, of oxidized ores from the Winnamuck mine, only sixty tons of outside ore (from the Spanish mine) having been smelted. The latter, like the principal Winnamuck ore, was oxidized or so-called carbonate ore. There was mixed with these oxidized ores three hundred to four hundred tons, or 7 to 10 per cent. of galena, some of which was mined with the oxidized ore, while a part was mined separately from the lower portion of the mine, and afterwards mixed with the ore, with a view of preventing the formation of deposits of metallic iron in the furnaces. The average assay in silver of all the ore handled was 51.46 oz. per ton, most of it existing as chloride of silver. The lead contents were 34.98 per cent,, all, or nearly all, in the form of carbonate of lead. The relative amount of silver is not at all constant, the best silver ore often being poorest in lead. The predominant gangue was silica, several determinations of which have been made on representative samples, yielding in three such samples 26, 38, and 58 per cent. silica, the latter test being ore containing but little lead. The average contents in silica are about 35 per cent., with 6 to 7 per cent. sesquioxide of iron, and small quantities of alumina and lime. Mechanically, the ore was very fine, and so thoroughly disintegrated that it presented few distinguishing characteristics, rendering sorting or separating of ore from waste difficult, and often impracticable. Experiments on a small scale have been tried, with a view to separate the silica by washing; but these were unsuccessful, as the finest slime, requiring a long time to settle in still water, contained a large amount of silver; and on a careful sizing and washing of the sands and coarser parts, the silver contents were found to be less dependent upon specific gravity

By Lawrence Addicks

SOME three years ago the smelter connected with the Chrome, N. J., refinery of the United States Metals Refining Co. found itself embarrassed y constantly increasing piles of unsmelted blast-furnace flue dust. The charge was fine and 10 to 15 per cent. of it was blown over into the flue. Sulphur was at that time too scarce to make the sintering of the finer part of the charge attractive, and raw-ore smelting in a reverberatory with a partly oxidized charge. was not to be thought of. A great many schemes were considered, including leaching, blowing into the converters, and briquetting, the problem finally narrowing down to blast roasting and nodulizing. It was found by experiment that blast roasting would not yield a satisfactory sinter without the addition of coke dust or sulphide fines to augment the fuel value of the flue dust, and on this account, together with the low first cost of a kiln, the latter was decided upon. A couple of tons of flue dust Were tried out in the experimental kiln in the laboratory at Yorktown, Va., of J. H. Payne, who acted as consulting engineer in this connection. This kiln was approximately 2 ft. in diameter and 20 ft. long and Was fired with fuel oil. A run of several hours yielded excellent nodules. The kiln showed no tendency to build in and form "nose rings" and gave a fuel consumption of 50 gal. of oil per ton of flue dust. There was no indication of any stack loss. The roasting was strongly oxidizing. A study of what full-size kilns were doing on cement clinker and on nodulizing pyrites cinder indicated that a 60 by 6 ft. kiln would certainly not consume over half of the oil per. ton of flue dust shown on the test and as fuel oil at that time Was selling around 2 1/2 c. per gallon it was decided to let it go at that.

Jan 7, 1914

By Frank C. Smith

The following article covers the history of a metallurgical campaign, commenced in March, 1905, at the mines of the Socorro Bold Co., in the so-called desert region of Yuma county, Arizona. The results obtained were simply a function of necessity, as will be seen from the description of the local economic conditions and the character of the ores treated. The mines and mill of the Socorro Gold Co. are located near the small village of Harrisburg, about 60 miles from Congress Junction, Ariz., which, at the time this work was undertaken, was the nearest railroad station, and from which all supplies were brought by wagon through the desert, at a cost varying from $10 to $15 per ton in addition to the railroad freights. Wood, the only available fuel, cost $7 per cord. Water for all purposes was pumped to the plant from a well in the valley about 4 miles away. Under the Arizona law, a working-shift is 8 hr. Wages were $6 per day for foremen and mill-men, $4 per day for blacksmiths and engineers, $3.50 per day for miners, and $3 per day for shovelers and trammers. The ore consisted of quartz, in which, above the 250-ft. level, the iron-minerale were largely oxidized and some free gold was visible; below that level few traces of oxidation occurred, and pyrite constituted the principal mineralizer in the quartz, together with occasional pockets of galena and a few eccentric specks of covellite. The 20-stamp mill was equipped with plates for amalgamation, and three Standard tables for concentration. The tails from the tables mere elevated by a centrifugal pump to a launder, by which they were conveyed to a 60-ton cyanide-mill for further treatment. The cyanide-mill was erected at so great a distance from the stamp-mill that a separate crew of workers and a separate power-plant were necessary. No arrangement had been made for the treatment of the concen-

Jan 1, 1907

By Michael Hoch, Daniel B. Butrymowicz

The isothermal section of the Ta-Ti-Zr-0 system at 1500°C was investigated using X-ray diffraction and rrzetallographic techniques. Up to 71.4 at, pct 0 the system contains nine four -phase regions. Zsopleths at 10, 20, 30, 40, 50, and 60 at. pct 0 were constructed. The purpose of this investigation was to determine the general shape of the quaternary equilibrium phase diagram of tantalum, titanium, zirconium, and oxygen at 1500°C. The system was truncated at 66.7 at. pct 0 on the Ti-0 and Zr-0 binaries and at 71.4 at. pct 0 on the Ta-0 binary. The Ti-Zr-0 system was investigated in this laboratory.' The binary systems of interest have been compiled and discussed by Hansen and Anderk, chmidt, Hayes, Roberson, and asche, and Hoch and athur.' EQUIPMENT The Same equipment as that for the study of the Nb(Cb)-Ti-Zr-0 system was used.' The X-ray dif- fraction patterns were analyzed with the help of the ASTM card set7 and NBS circulars.' MATERIALS The tantalum powder (99 pct pure), the titanium powder (99.6 pct pure), the tantalum pentoxide, and the zirconium dioxide used in this study were purchased from the Fairmont Chemical Co., Newark, N.J. The zirconium powder (99.4 pct pure) was obtained from the Charles Hardy Co., Inc., New York. Reagent-grade titanium dioxide was purchased from the Matheson Co., Inc., Norwood, Ohio. The oxides were dried in air at 700°C for 24 hr before use. Though the materials used were not "hyper-pure", the impurities present do not affect the results (lattice parameters, phase boundaries), within the experimental accuracy. PROCEDURE The samples were heated by induction in a vacuum furnace, and examined metallographically and by X-ray diffraction at room temperature. The procedure was the same as that in the Nb(Cb)-Ti-Zr-O work,' EXPERIMENTAL RESULTS AND DISCUSSION OF DATA A four-component system can be represented by a tetrahedron at a single temperature and pressure

Jan 1, 1964

By C. D. Hunter, I. B. Browning, N. W. Shiarella

The oil industry in Kentucky during the year 1937 shows some improvement over that of 1936, although not to as marked a degree as 1936 over 1935. Oil in Eastern Kentucky In eastern Kentucky more than 150 new inside wells were completed in the pools of Estill, Powell and Lee Counties. Also, a number of wells were completed to be used as intake wells in repressuring operations. Repressuring is proving more successful in this area than was expected, as the producing formation was the Corniferous limestone. However, when the repressuring work was started, cores taken of this formation showed that the oil in the second and third pays was from a true sand, only the first pay in the top being a sandy lime formation. In the Weir sand fields of Magoffin, Johnson and Lawrence Counties, continued repressuring operations caused an increase in the quantity of oil. These operations and the completion of 20 new wells in the Maxon sand and the Big lime (Mississippian) in Martin and Floyd Counties produced an increase in the oil production from eastern Kentucky of 100,000 bbl. over that of 1936. As an illustration of what may yet be done in the fields of eastern Kentucky, one repressuring operation involving four oil wells and one pressure well produced an increase in production, or an additional recovery of 7200 bbl. of oil during the year. This amounts to a recovery of 2000 per acre for the 3.6 acres involved. In other fields of eastern Kentucky, as in the Corniferous pools of Estill, Powell and Lee Counties, increases in the runs are noted as follows: In the area of the Estill County pool from which the oil passes through the Revena pumping station, the production in 1935 was 88,175 bbl.; in 1936, was 102,849 bbl.; and in 1937 was 107,704 bbl. This pool was

Jan 1, 1938

By Linn Bradley

F. H. WILLCOX, Pittsburgh, Pa. (communication to the Secretary*). -We must keep in mind, in balancing the savings-to be anticipated by the most efficient combustion of gas, the best heat absorption by stoves or boilers, and by heat conservation of the thermal equivalent carried in. blast-furnace gas-against the cost in operation, maintenance, depreciation and interest charges of the equipment required to realize the savings, that in boilers and hot-blast stoves we are basing our economies upon two different, combustion phenomena. In a boiler, evaporating water at a heat efficiency of 70 per cent., approximately 47,550 B.t.u. is consumed per boiler horsepower. Taking the blast, furnace gas cited in the above paper, we are using about 38 ¾ lb. of cold-washed gas per boiler horsepower. Using the hot-cleaned gas, the same equivalent heat is furnished by 34 1b. of the gas , This leaves of the 38 ¾ lb. required in the case of cold-washed gas, 4 ¾ lb. or 6,650 B.t.u. available for raising more steam when dry-hot gas is used, which means that an equivalent amount of coal firing can be dispensed with. Per ton of iron, this figure 2/3 X 12,000/38.75 X 4.75 = 980 lb. gas or 1,372,000 B.t.u. or 100 lb. coal. At $1.50 per ton of coal, the net saving per ton of pig iron produced by the use of dry-hot over wet-cold cleaning will, therefore, be about $0.06 2/3. At a stove heating 8,500 lb. of air per toil iron from 150° F., one-quarter of the dry-hot cleaned gas, or 3,000 lb., would heat the air to 1,450° F. 3,000 X 1,400 X 0.70 = (8,500 X 0.255) = 1,300. The cold-wet gas would heat the same amount of air to 1,280° F. 3,000 X 1,225 X 0.70 = (8,500 X.0.255) = 1,130.

Jan 4, 1917

By W. L. DeMoulin

A rather detailed description of the entire plant and leaching process was given in a paper recently presented to the Institute,I so this paper will cover briefly only the crushing practice of the New Cornelia Copper Co. for the year 1918. The ore, which is mined by steam shovels and loaded in side-dump cars, passes through two crushing plants. These crushing plants were constructed to crush the ore required by a leaching plant of 5000 tons daily capacity, during a crushing period of 16 hours. The primary plant reduces to less than 3 in. (76 mm.) and the secondary plant, to 1/4 in. (6.35 mm.), which is sufficiently fine for satisfactory percolation of solution through the ore in the leaching tanks, and to give-good extraction. The ore is then conveyed to the leaching tanks by a system of 28-in. (71-cm.) belt conveyors. On the way to the leaching tanks, it passes through an automatic sampling plant, where a 1-per cent. sample is taken. From the primary crushing plant, the ore is conveyed by two 36-in. (91-cm.) belt conveyors to a 10,000-ton steel storage bin, with a reinforced-concrete flat bottom. The ore discharges from the bottom of this' bin onto four 20-in. belt conveyors, which deliver it to the four units of crushers in the secondary crushing plant. There is no storage bin between the mine and the primary crushing plant. The ore breaks very coarse, is very hard, and contains a great many boulders, some as large as 4 1/2 by 4 1/2 by 10 ft. (1.3 by 1.3 by 3 m.). Jams form in the bowl of the coarse crusher in the primary crushing plant, as a consequence, and are freed by means of an immense steel hook operated from a 40-ton electric traveling crane. Primary Crushing Plant The primary plant consists of one No. 24 Gates, style K, gyratory crusher, followed by four Gates No. 8, style K, gyratory crushers. All

Jan 1, 1920

By Frank A. Wilder

The most important process in a gypsum mill is calcining the crude mineral. There seems, however, to be little progress or change in calcining methods. This would not be surprising if the industry was not, in other ways, developing. The demand for gypsum in its various forms, however, is growing rapidly. The output has more than doubled in six years, while the value of gypsum products, through the introduction of higher priced. specialties like wall board, has nearly trebled in the same period. In other industries, under somewhat similar conditions, calcining in rotating cylinders, either by direct or indirect heat, has replaced the method that still prevails in gypsum mills. Efforts were made to use rotary calciners in manufacturing plaster of Paris in the early days of the industry in America; 25 years ago rotary calciners were in use in Germany. At that time there was great diversity of opinion in regard to their merit. In 1923, the output of calcined gypsum in the United States was 3,101,378 tons. About nine-tenths of this material was calcined in kettles, the remainder in rotary calciners. The ratio was about the same 10 years ago. During these 10 years, rotary calcining made relatively no gains, and there were no important changes in the construction of kettles. In their manipulation, the only change of consequence was in connection with the length of time that the gypsum is left in the kettle. For wall-plaster purposes, all producers who had not previously done so changed from second-settled to first-settled calcined gypsum, thereby reducing production costs and improving the plasticity and sand-carrying capacity of their plaster. Nevertheless, in the few mills in which they were installed, rotary calciners continue to operate, with possibly one or two exceptions, and to be highly successful. In these 10 years, the Cummer type of rotary calciner has disappeared. All rotaries now used are of the cement-mill type with a diameter of 6 or 7 ft. and length of 80 to 120 feet.

Jan 1, 1925

By T. T. Quirke, L. C. McCabe

The object of the present investigation was to discover a physical basis for rank differentiation of coals, particularly the coals of the Illinois basin. Vitrainl was selected as the most appropriate coal material to use in the investigation, because of its homogeneity and ease of acquisition and handling. Previous Investigations.—Hoffman and Jenker2 have shown that their index of reflection is a characteristic of coal rank and that the Leitz slit microphotometer may be used in determining the different intensities of reflection. Zhemchuzhnikov3 measured the luster of several Russian coals with a color analyzer and found that the degree of luster was indicative of rank. Luster-volatile matter curves based on data from Sucan and Donetz coals were satisfactorily used for determining the rank or volatile content of other coals after the luster strength was determined. In 1934 Fisher4 reported refractive indices of 1.76 for bituminous vitrain and 1.87 for semibituminous vitrain. Investigation of Angle of Polarization.—Early in the present investigation several determinations of the refractive indices of the vitrains were made by immersing the minus 150 plus 200-mesh powder in oils. It was found that all the samples fell within the range 1.6 to 1.8. However, difficulty was encountered in securing oils in this high index range that were closely enough spaced to separate some of the coals. The writers made a number of determinations of the angle of maximum polarization from the polished surfaces of the vitrains and derived their refractive indices by the procedure described in the following paragraphs. Johannsen5 states the fundamental concepts underlying this procedure as follows: In general, in an isotropic medium light vibrates in all directions at right angles to the direction of propagation, and when it meets with another isotropic medium at an angle, part of the light is reflected and part refracted. It has been found that after reflection or refraction, the vibrations do not move with the same freedom in every

Jan 1, 1938

By T. T. Quirke, L. C. McCabe

The object of the present investigation was to discover a physical basis for rank differentiation of coals, particularly the coals of the Illinois basin. Vitrainl was selected as the most appropriate coal material to use in the investigation, because of its homogeneity and ease of acquisition and handling. Previous Investigations.—Hoffman and Jenker2 have shown that their index of reflection is a characteristic of coal rank and that the Leitz slit microphotometer may be used in determining the different intensities of reflection. Zhemchuzhnikov3 measured the luster of several Russian coals with a color analyzer and found that the degree of luster was indicative of rank. Luster-volatile matter curves based on data from Sucan and Donetz coals were satisfactorily used for determining the rank or volatile content of other coals after the luster strength was determined. In 1934 Fisher4 reported refractive indices of 1.76 for bituminous vitrain and 1.87 for semibituminous vitrain. Investigation of Angle of Polarization.—Early in the present investigation several determinations of the refractive indices of the vitrains were made by immersing the minus 150 plus 200-mesh powder in oils. It was found that all the samples fell within the range 1.6 to 1.8. However, difficulty was encountered in securing oils in this high index range that were closely enough spaced to separate some of the coals. The writers made a number of determinations of the angle of maximum polarization from the polished surfaces of the vitrains and derived their refractive indices by the procedure described in the following paragraphs. Johannsen5 states the fundamental concepts underlying this procedure as follows: In general, in an isotropic medium light vibrates in all directions at right angles to the direction of propagation, and when it meets with another isotropic medium at an angle, part of the light is reflected and part refracted. It has been found that after reflection or refraction, the vibrations do not move with the same freedom in every

Jan 1, 1938

By J. E. Lawver, R. R. Beebe, J. L. Wright

This paper describes a separation coefficient Ks useful as a "measure of goodness" to be used in factorial and response surface methodology. The use and limitations of the separation coefficient is discussed, and a computer program for calculating the coefficient is presented (in Fortran). The application of the coefficient is illustrated in terms of electro-dynamic separation of Mesabi semitaconites. Present Mesabi wash ores and gravity concentrates with an iron to silica ratio of only 5 or 6 to 1 are no longer readily salable, for the steel industry now demands concentrates with an iron to silica ratio of at least 10 to 1 and a structure comparable to pellets. THE ELECTRODYNAMIC-FLOTATION PROCESS Wash ores and gravity concentrates consist largely of hematite, goethite, and quartz. Several methods of beneficiation are being tested at the University of Minnesota Mines Experiment Station with a view to improving the salability of these materials. A process that looks particularly promising from an economic view is the combination electrodynamic-flotation process shown in Fig. 1. This process is designed to upgrade gravity concentrates to a 5% silica content and to provide fines — from the flotation circuit —to permit agglomeration. Although electrodynamic concentration of specularite is an established and well-understood process, the electrodynamic concentration of Mesabi goethite-hematite gravity concentrates presents interesting and enigmatic problems. The gravity concentrates are by no means assemblages of clean, free mineral particles. Instead, the concentrates to be retreated consist of a great variety of middlings. More or less complete liberation would require grinding to minus 325 mesh, at which size only flotation could be expected to function, and even then grades and especially recoveries would not be satisfactory. A close examination of the gravity concentrates shows that grinding to 100% minus 48 mesh will produce a product which is approximately 70 weight % plus 325 mesh (see Fig. 1). The remaining 30 weight % can be deslimed and floated by conventional methods, but the resulting iron concentrate is

Jan 1, 1964

By Oswald J. Heinrich

The serious objections to the Leblanc soda process may be enumerated as follows: 1st. The total loss of sulphur employed, equal to about one-third of soda produced. Various processes have been proposed to regain this sulphur, but none have come into practical operation on a large scale. 2d. The great balk of waste products, equal to nearly twice the amount of soda, which forms a serious annoyance about the works. 3d. The complicated series of operations necessary. 4th. The large amount of labor and fuel required. 5th. The prodoction of noxious gases which necessitate extensive condensing arrangements. 6th. The large capital necessary for the . erection and maintenance of the plant, These objections have long been recognized, and many attempts have been made during the last thirty or forty years, to discover and introduce a cheaper and more direct method for the manufacture of this important staple. The reaction between bicarbonate of ammonia and common salt, whereby bicarbonate of soda, insoluble in a solution of chloride of ammonium, is precipitated, has been closely studied for the last twentyfour years, on account of its simplicity and the purity of the product it affords. This ammonia-soda process is now so far developed, that while it is not in position to entirely replace the Leblanc process, yet it can successfully compete with it where the conditions and localities are favorable. The advantages of the ammonia-soda process may be summed up as follows : 1st. It is independent of the essential agent in the Leblanc process, via., sulphuric acid, and hence can be established in those localities where salt is abundant, and yet where sulphuric acid could not be made, or where it could not be transported profitably. 2d. The ammonia-soda process being a " wet" process, the evaporation of salt brines to obtain the solid salt required in the Leblanc process, is rendered unnecessary. 3d. The direct conversion of chloride of sodium into bicarbonate of soda, which can readily be converted into the maketable monocar-

Jan 1, 1879

By R. N. Lambert

This chapter discusses the evolution of high-speed shaft sinking in the South African goldfields. Whether we are talking of an additional shaft at an existing mine or the opening of a new mine, the increasing depth of the ore-bearing reef and the unremitting inflation of costs have provided the compelling incentive to shorten the time taken to sink a shaft and start earning revenue. The high costs of sinking a modern shaft system in South Africa are indicated by the recent example of a deep-level mine in the Orange Free State where the cost was nearly $22!42 million. This was for a single large shaft of 33 ft 6 in. diam (in all cases, shaft diameters refer to the finished diameter of the shaft, e.g., inside the concrete) and 7600 ft deep with a prestressed concrete brattice wall dividing the shaft into upcast and downcast compartments. Higher costs and the consequent demand for increased tonnages, deeper shafts with higher rock temperatures, which necessitate highly sophisticaled ventilation and refrigeration requirements, have all accentuated the importance of time in a shaft-sinking operation. Rectangular shafts with their simple design and comparatively reasonable sinking speeds remained popular for many years until greater mining depths combined with a host of the other allied problems inevitably led to changes in shaft design. Circular shafts appeared to offer many advantages, particularly when considering ventilation, strength, mechanization, and sinking through bad ground and heavily watered zones. However, these advantages were largely offset by the slow rate of advance necessitated by the need to stop sinking in order to line the shaft. The development of the curb ring which enabled lining to take place above the bottom of the shaft and the use of multideck stages allowing sinking and lining to be carried out concurrently resulted in a major break- through. Rectangular shafts, which are still popular elsewhere, have been superseded in South Africa by circular shafts which are almost without exception the only shafts sunk in our goldfields today.

Jan 1, 1970