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The Library of the above-named Societies is open from 9 A.M. to 10 P.M., except on holidays. It contains about 70,000 volumes and 90,000 pamphlets, including sets of technical periodicals and publications of scientific and technical societies. Members of the Institute, with few exceptions, are forced to spend a portion of their time in localities isolated from sources of information. To these the Library, through its, Library Service Bureau, can render valuable service through correspondence; letters requesting information will receive especial attention. The Library is prepared to furnish references and photographic copies of articles on mining and metallurgical subjects; to determine the existence of mining maps, and to furnish general information on the geology and mineral resources of all countries. All communications should be made as definite as possible so that the information received may be what is desired and not include collateral matter which may not be of interest. The time spent in searching for such collateral matter will be saved, and the information will be sent more promptly anti in more usable shape.

Jan 1, 1918

By Joseph Taylor

THE Ducktown district is in the extreme southeastern corner of Tennessee, its principal railroad point being Copperhill, on the Blue Ridge division of the Louisville &Nashville Railroad, midway between Knoxville, Tenn., and Atlanta, Ga. The ore deposits were discovered about 1849 and development work was begun in the early fifties. Only the "black copper? ores, products of secondary enrichment, from beneath the gossan outcrops and overlying the massive sulphide deposits, were then mined and smelted for their copper content. Upon their exhaustion, attention was directed to the treatment of the pyritic ore remaining, resulting in the devising of a, process consisting of open-heap roasting, followed by blast-furnace smelting, which was practised until 1903, when semi-pyritic smelting was-adopted. In 1908, the companies operating in the district; on account of litigation brought about by alleged smoke damage to forests and vegetation in the adjoining counties of Georgia, were compelled to erect enormous acid plants to prevent the escape of the sulphur dioxide produced in the smelting operations. The utilization of this sulphur dioxide in the, manufacture of sulphuric acid as a byproduct has added to the revenues of the companies, and at the same time has furnished a valuable source of supply of sulphuric acid to the fertilizer trade of the South, and more recently to the manufacturers of explosives for use in the production of war munitions. The steadily increasing demand for sulphuric acid for the manufacture of fertilizers and explosives, together with the decreasing imports of pyrites, from which the acid is generally produced, now makes it imperative that greatly extended resources of pyrite be developed. While the Ducktown deposits do not predominate in pyrite, they do contain an inexhaustible supply of pyrrhotite, admixed with pyrite, which can be satisfactorily employed to supplement or substitute for pyrite in acid making.

Jan 2, 1918

By Clyde Ross

The Thunder Mountain mining district is in eastern Valley County, Idaho. It is of undefined extent but nearly all of the development is confined to the ridge between Monumental and Marble creeks in unsurveyed T. 19 N., R. 11 E. (Fig. 1.) The district is still remembered among the mining men of Idaho because of the ill-advised rush there in 1902, but available evidence indicates that optimism regard-ing the future development of the gold deposits of the district is warranted. The former disappointment arose from the fact that many who took part in the early rush hoped for bonanza deposits, whereas the district is characterized by large masses of low tenor. In 1926 the writer,, assisted by C. Y. Behre, 3r., incidental to other work in the region, devoted four days to the detailed examination of the then access-ible workings in the district. Tha major workings in the Sunnyside, Venable, Standard and H-Y mines, most of which were still in good shape, were examined., Surface exposures at the Dewey were studied, but the main workings there have long been caved. The time available did not permit more than a cursory examination of the surface at a distance from the mines. Sufficient data have been gathered, how-ever, from published and unpublished accounts, production records and mine maps so that the major features of this deposit, by far the most productive in the dis-trict, can be summarized. The kindness of Messrs. Dan McRae and R: A. Davis in making available information regarding the mines of the district, a rd of these men and others in recounting early history and their observations in now inaccessible workings, have aided materially in preparing the present paper. A brief summary of the salient characteristics of the district has already been published.1/ Cir-cumstances have contributed to delay in the publication of the full official report, which will contain mine maps and ?such other detailed data as could be ontained. In view of the present interest in gold deposits it seems desirable to present here a summary of results of the study.

Jan 1, 1933

By Evangelo Agelopas

Steelmaking operators are familiar with the costs involved when ingot molds must be scrapped prematurely because of the failure of a stripping or turning lug. Some mold sizes can be expected to produce one hundred or more ingots per mold before normal wear causes the mold to become unserviceable. Therefore, the failure of lugs at less than the average life for that mold size represents a sharp increase in terms of mold cost per ingot ton produced. To combat these wasteful and costly occurrences, it was determined that an efficient method of replacing lugs should be developed to extend the life of molds which are otherwise in good condition. The purpose of this paper is to describe a lug welding method which has proved effective and economical. This method is known as the Oxygen and Mapp Process.

Jan 1, 1972

By William M. Shakespeare

Thin sheet copper in wide widths and long lengths is being made electrolytically on a commercial scale at the Raritan Copper Works, Perth Amboy, N. J. This achievement is the result of three years of intensive development preceded by almost ten years of intermittent research. While the idea of producing sheet copper by the electrolytic process is old, the literature on the subject is limited. It is evident, however, that the idea has been attractive to many investigators for years, if the issued patents in the field may be taken as an indication. These patents may be divided into three groups. In one group metal is deposited on some form of endless cathode, of which a revolving cylinder or drum is the most common. The cathode moves slowly and at such a rate that metal of the desired thickness is deposited during one pass through the electrolyte. The deposited metal is stripped off the cathode at a point above the solution level and, after washing and drying, is wound up immediately. In the next group of patents a band of thin metal is increased in thickness by being guided through an electrolytic tank in opposition to suitable anodes. The third group covers means of producing sheets of definite dimensions where flat plates or revolving cylinders are used as cathodes. With such cathodes, sheets of definite dimensions are built up to the desired thickness and then stripped off. The process as worked out at the Raritan Copper Works combines the methods of the first two groups. In the first stage a sheet weighing 1 oz. to the square foot is produced on a revolving cylinder. This sheet is increased to any desired thickness in the second stage. In the first stage (Fig. 1) a lead-covered copper drum is mounted horizontally on a lead-lined concrete tank so that it. is partly immersed in the electrolyte. The cylindrical surface of the drum immersed in the electrolyte is opposed by lead anodes spaced about 36 in. from it. The electrolyte, made up of appropriate quantities of copper and sulfuric acid, flows into and out of the tank continuously. The ebullition of gas at the surface of the anodes induces a vigorous circulation of electrolyte between the anodes and the drum surface, which is augmented by means of compressed air introduced through a pipe between the ends of the anodes under the center of the drum, not indicated on the diagram. This

Jan 1, 1934

By J. T. Norton, A. L. Mowry

The monocarbides of the A subgroup elements in the fourth and fifth group of the periodic table in addition to being hard and refractory are of special interest in that they are isomorphous in crystalline structure. They are cubic with a sodium chloride type structure in which the metal atoms are essentially close packed in a face-centered cubic arrangement with the carbon atoms placed in the interstices between. Interstitial structures of this close packed type were first investigated systematically by Egg1 and he gave the rule for their formation, stating that the radius ratio of the nonmetal to the metal atom should not exceed the value of 0.59. The carbides of interest are those of titanium and zirconium of the fourth group and vanadium, columbium and tantalum of the fifth group. Table 1 shows the radius ratio using the Goldschmidt radii for 12 coordination for the metal atoms and the diamond radius for the carbon atom. It will be noted that while there is considerable variation in the size of the metal atom, in all cases the ratio is smaller than the limit of 0.59 placed by Hägg. It has been known for some time that these cubic carbides are soluble in one another, at least to some extent or, in other words, the metal atoms can be replaced, one by another without destroying the stability of the structure. Since the stability of these close packed interstitial substances appears to depend more upon geometry than upon the exact chemical nature of the atoms involved, it is of interest to examine the possibilities of replacement in these carbides in some detail. Hume-Rothery2 has pointed out the importance of the difference in size of solute and solvent atom as a factor in limiting the solubility in simple binary solid solutions. Largely on an empirical basis, he states that if the difference in size between solvent and solute atom is more than 14-15 pct of the solvent atom, the range of solubility is very restricted. The atom size was based on the distance of closest approach in the elements involved. While there is some question as to how one should calculate the size of the metal atom in the carbide structures, reference to Table 1 will show that zirconium is the largest and vanadium the smallest of the group and that the difference is about 15 pct. The Ti-Zr difference is about 9 pct and the others are smaller. Thus one would predict that if the size factor controls the solubility, all of the pairs except VC-ZrC would have wide or complete solubility whereas this latter pair is on the border line and might have restricted solubility. The purpose of the present investi- gation was to examine the solubility of the several pairs of carbides by heating them together until equilibrium was established and then examining the product by X rays. Previous Work Agte3 and his associates prepared various transition metal carbides and determined the melting points of binary mixtures. He concluded from the shapes of the melting point curves that there was extensive solubility in the case of the cubic carbides. Umanskii and his colleagues made an investigation of a number of pairs of the cubic carbides, using X rays and plotted lattice parameter vs. composition curves for the systems TaC-Tic, CbC-Tic, TaC-ZrC and CbC-ZrC. All pairs showed a continuous series of solid solutions. The first two pairs gave a linear relation while the latter two showed a negative deviation from Vegard's law. Kiefer and Nowotny, in a paper which became available after the present work was well advanced, investigated the binary pairs of the five cubic carbides by means of X rays. Relatively few points were obtained and results indicated that in some cases, at least, equilibrium was not reached at the temperatures used. The results indicated that solubility in the VC-ZrC system was not complete. All of the results of previous investigations indicated the desirability of a more detailed study. Materials The raw materials used were mono-carbides of titanium, zirconium, vanadium, columbium and tantalum and were the purest which could readily be obtained commercially. Spectrographic qualitative analysis showed that the CbC and TaC contained less than 1 pct

Jan 1, 1950

By Gilbert Thiessen

The use of a pneumatic process for removing fine material (usually less than 48-mesh) is increasing in the preparation of stoker-size coal. The fine material, produced as a byproduct, is known as dedusting-plant dust. Since fusain is the most friable ingredient in bituminous coal (it may readily be rubbed to an impalpable powder with the fingers), it tends to accumulate in this byproduct. While studying the properties of cokes produced by blending Illinois coals with other materials, considerable quantities of fusain were required. Dedusting-plant dust suggested itself as a convenient supply. It became necessary, therefore, to determine how much fusain was contained in a measured sample of dust, how the fusain was distributed with respect to size and specific gravity in the sample, and whether or not it would be feasible or desirable further to concentrate or to recover the fusain from the dust before using it in the blends to be coked. The results obtained indicate the possibilities of production of a high-fusain material. Previous Work Lehmann and Stach have developed a processl,2 for separation of banded bituminous coals into their band ingredients by impact pulverization followed by screening or air separation. They found, as might be expected, that the fusain was concentrated in the very fine material. Prockat and Raute3 examined samples from dedusting plants operating in the Ruhr. They separated the dust into various screen sizes and determined the composition of each of these sizes in terms of the banded ingredients of the original coal. In the dusts examined by them, the finest material was always high in fusain, and most of the fusain contained in the original samples could be recovered from this fraction.

Jan 1, 1936

By Gilbert Thiessen

The use of a pneumatic process for removing fine material (usually less than 48-mesh) is increasing in the preparation of stoker-size coal. The fine material, produced as a byproduct, is known as dedusting-plant dust. Since fusain is the most friable ingredient in bituminous coal (it may readily be rubbed to an impalpable powder with the fingers), it tends to accumulate in this byproduct. While studying the properties of cokes produced by blending Illinois coals with other materials, considerable quantities of fusain were required. Dedusting-plant dust suggested itself as a convenient supply. It became necessary, therefore, to determine how much fusain was contained in a measured sample of dust, how the fusain was distributed with respect to size and specific gravity in the sample, and whether or not it would be feasible or desirable further to concentrate or to recover the fusain from the dust before using it in the blends to be coked. The results obtained indicate the possibilities of production of a high-fusain material. Previous Work Lehmann and Stach have developed a processl,2 for separation of banded bituminous coals into their band ingredients by impact pulverization followed by screening or air separation. They found, as might be expected, that the fusain was concentrated in the very fine material. Prockat and Raute3 examined samples from dedusting plants operating in the Ruhr. They separated the dust into various screen sizes and determined the composition of each of these sizes in terms of the banded ingredients of the original coal. In the dusts examined by them, the finest material was always high in fusain, and most of the fusain contained in the original samples could be recovered from this fraction.

Jan 1, 1936

By A. W. Hahn

ALTHOUGH enormous tonnages of sulfide lead and lead-silver ores are treated by flotation, the products of flotation mills treating oxidized ores of lead and silver are almost negligible. However, each year there is a slight tonnage increase, and it is possible, should a sizeable oxidized lead ore tonnage be developed, the field will become an important one. In 1918, the Shattuck-Arizona Copper Co. erected a plant at Warren, Ariz., for the treatment of oxidized gold-silver-lead ores. Glenn L. Allen worked out the metallurgy for this plant, employing tabling, followed by flotation concentration. The Shattuck-Denn Mining Co., successors to the Shattuck-Arizona Co., now employs straight flotation in the same mill, treating about 200 tons daily. In 1922, the San Diego mill of the American Smelting & Refining Co., in the Parral district, Chihuahua, Mexico, started operating a 600-ton mill on a mixed sulfide-oxide gold-silver-lead ore. Tables and flotation are used here, and about 200 tons of oxidized slime treated by flotation. In 1923, the Eureka Holly Mining Co. started operating a 90-ton mill at Eureka, Nev., using tables and flotation, on an oxidized lead-silver ore. In 1924, the American Smelting & Refining Co. started a 250-ton mill in northern Peru, to treat a high grade oxidized gold-silver ore with very low lead values. In 1924, the Grand Central Mining Co. started a 200-ton flotation mill on old oxidized gold-silver-lead tailings at Fairbank, Ariz. In 1926, the Compania International Minera, a subsidiary of the Union Corporation of London, started a 250-ton flotation mill on oxidized lead-silver ores, and during April of this year, the Phelps-Dodge Corp. will probably start a 250-ton mill at Warren, Ariz., on oxidized lead-silver ores, similar to those being treated in the Shattuck-Denn mill.

Jan 1, 1927

By Donald Barton

THE elastic, earthwaves produced naturally by earthquakes -have been used for a long time as evidence from which to draw conclusions in re-gard to the constitution of the interior and crust of the earth. The elastic earthwaves produced artificially by occasional explosions and recorded by seismological observatories were used to draw conclusions regarding the-geology of the intervening area. Intentionally pro-duced elastic earthwaves in the past five years have come to be used extensively in the investigation of very local geologic structure. The seismic method of work-ing geologic structure, which makes use of artificially controlled explosions and the resulting earthwaves, has developed almost `clairvoyant power in handling certain geologic situations, and in the discovery of salt domes in the Gulf Coastal: Plain region of Texas and Louisiana has 'scored the, most brilliant success. The. first, proposal for the use of artificial earth-quakes in the study-of velocity of elastic earthwaves in the surface formations of the earth's crust was made before 1888 by the English seismologists, Mallet and Abbot. Partly on the basis of their work, A. Schmidt in 1888 proposed- the use of the time-distance graphs of artificial earthquakes to study the variation of ve-locity with depth. Belar in 1902 proposed the practical application of such investigation in connection with boring tunnels. Galitzen repeatedly (1912, 1913) pro-posed the use of explosions to study the velocity of the longitudinal and transverse waves in the uppermost formations and pointed out that the velocity depended in a high degree on the physical character of the beds and that from changes in the velocity, conclusions could be drawn in regard to the compositions of the beds. Somewhat the same thought: was proposed by von dem Borne (1908), by Benndorf, Udden, and others.

Jan 9, 1928

By W. C. Ralston

This note will give the cost of driving an adit at the Melones mine, in 1898, and, for purposes of comparison, the cost of similar work, in 1888, at the Hogsback mine, Placer county, Cal. The property of the Melones Consolidated Mining Company is situated on the Mother Lode of California, on Carson Hill, near Robinson's Ferry, and adjoins on the south the famous Morgan mine. The Mother Lode of California extends for a distance of over 100 miles in length through several counties, with a general strike of northwest and southeast. Three veins (commonly called the east vein, middle vein and west vein) comprise the " lode " in this section. The Melones property is situated upon the east and middle vein, and includes 1407 feet on the east vein and 5166 feet upon the middle vein. A shaft had been sunk on the Reserve mine, which is on the east rein, and within 230 feet of the Morgan mine. At the bottom of this shaft drifts and cross-cuts had been run and a large body of ore developed. It was deemed advisable by the Melones Mining Co., which holds this property under bond from the Melones Consolidated Mining Co., to run a tunnel and develop this body of ore at a greater depth. An option was obtained upon the South Carolina mine, which adjoins the property on the south, and is on the same east vein, upon which a tunnel had been run a distance of 1080 feet. This tunnel was cleaned out and continued to a point within forty feet of the Morgan mine. It passes directly under the Reserve shaft, 425 feet below the 200-foot level. The additional length of tunnel to be run was 1923 feet, and

Jan 1, 1899

By Robert L. Davidoff, Rodney D. Rosenkranz

In order to determine copper resource data and production costs for major market economy mines and deposits, the Bureau of Mines has performed detailed engineering and economic analyses on 271 of the world's most significant copper deposits. At the demonstrated resource level, copper potentially recoverable from these deposits is estimated to be 325 million metric tons, 180 million tons of which would be from Pacific Rim countries. Chile has the largest known copper resources of any country and accounts for nearly 45 percent of the copper avail- able from Pacific Rim countries. At $1.00 per pound, in January 1981 dollars, an estimated 80 million metric tons of copper are potentially economically recoverable from Pacific Rim deposits. At this same price, analyses indicate that these deposits could economically produce about 2 million metric tons of copper per year. Sensitivity analyses reveal that if operating costs were to decrease 25 percent, with copper at a price of $1.00 per pound, total copper economically recoverable from Pacific Rim countries could increase by about 10 percent; a 25 percent increase in operating costs could lower economically recoverable copper by 50 percent. Analyses also indicate that byproduct prices are important. If byproduct revenues increased 25 percent, at a copper price of $1.00 per pound, an additional 2 million metric tons of copper could be recovered; if byproduct revenues decrease 25 percent, about 5 million metric tons less of copper metal would be available. Since the January 1981 date of this study, inflation and decreases in copper and byproduct prices have greatly affected the profitability of the world copper industry. Many producing mines have curtailed or suspended production, and development of many non- producing deposits has been postponed.

Jan 1, 1982

By Kim C. Harden

INTRODUCTION The purpose of the following discussion is to present the state of the art of solution mining. Since the economics of a mining method ultimately determines its applicability and viability this presentation shall revolve around the costs of in- situ solution mining. First the assumed physical characteristics of the hypothetical ore body are described, followed by the appropriate operating assumptions. Then after a brief discussion on the type of surface plant to be used, the assumed project time tables and costs for Texas and Wyoming are presented. Finally, the economics of in-situ uranium leaching are analyzed through the use of discounted cash flow rate of return analysis. ORE BODY CHARACTERISTICS The assumption of the ore body characteristics is probably the most variable portion of this discussion. The characteristics that have been used are based mainly on state of the art technology, however, consideration of the most common depths of ore, ore thicknesses, and permeabilities also influenced these assumptions. In addition, it is assumed that these assumptions are equally applicable to Texas and Wyoming. The average grade of the ore is assumed to be .09% U308 with no apparent disequilibrium. The average thickness of ore is 2.29 m (7.5 ft) which results in an average grade-thickness (GT) of .675. The assumed depth to the top of the ore is 121.92 m (400 ft), the ore density is placed at 1.78 gm/cc (18 cu ft/ton), the porosity is estimated to be 28% and the permeability 1 darcy. These assumed ore body characteristics are shown in Table I. In addition, it is specified that the costs to be later discussed are based on a minimum GT cut-off of 0.15. It is more common to use GT cut-offs of 0.30 to 0.50 but GT cut-offs as low as 0.15 in conjunction with a minimum grade of 0.05% U308 have been used in the past with success and is considered state of the art. The ultimate percentage of uranium recovered from the ore is left to the discretion of the reader since the costs and economics are based on pounds recovered by the surface plant. OPERATING AS.SUMPTIONS An annual production rate of 200,000 lbs U308!yr was chosen for this example. In order to maintain this production rate, based on the ore body characterized above, a flow of 4731 liter/min (1250 GPM) with a recovery solution grade averaging .039 gm U308/liter is assumed. A regular 5 spot well field pattern is used with a well spacing of 21.5 m (70.7 ft) between like wells and 15.24 m (50 ft) between unlike wells. This well spacing gives each well an area of influence equal to 232.25 sq m (2500 sq ftl. An excess wells factor of 1.17 is used to estimate additional monitor wells and well field boundary wells. Each production well is expected to yield an average flow rate of 37.85 liter/min (10 GPM). In addition it is assumed that the ore body has a good shape in that it is not tenuous and narrow but has at least an average width of 200 ft. The process chemistry required for this ore body is assumed to be based on the sodium carbonate System- Oxygen is the chosen oxidant. Sodium chloride elution followed by precipitation with hydrogen peroxide makes up the remaining portion of the process. A fluidized up-flow ion exchange system is specified. The operating assumptions are listed in Table II. Restoration of the ore body shall be assumed to be accomplished through the use of ground water flush. Other methods may be considered as having to fall within the costs estimated for a ground water flush in order to be economic. In Texas it is assumed that a high capacity disposal well (200 GPM +I is required and in Wyoming evaporation ponds covering approximately 35 acres are to be used. No specific cost has been given to restoration. Instead only the additional capital investment for restoration equipment is given. Then, one year of restoration operating expense is estimated and included as the operating expense for one year beyond the last pound of U308 produced. It is also assumed that restoration will be pursued in the mined out areas of the ore body contiguous with ongoing production.

Jan 1, 1980

By J. H. Hamilton, J. R. Lewis, J. C. Nixon, C. L. Graverson

Recently there has been much speculation concerning the advantages of using oxygen enriched air or pure oxygen in pyrometallurgical processes. The advantage of using oxygen in the iron blast furnace and elsewhere in ferrous production metallurgy has been given careful consideration and, as a result, several experiments have been tried and a number of papers have recently appeared on this subject.1,2 The use of oxygen in nonferrous smelting likewise poses interesting problems, therefore any experimental data obtained in this area should likewise be of general interest.3 Ashcroft4 in 1933 claimed to have obtained improved results when copper concentrates were roasted in pure oxygen. However, his paper gives neither data nor experimental details. A few years ago a research program was started in this laboratory which had for its objective the determination of the mechanism of oxidation of metallic sulphides under varying conditions. To date we have studied the oxidation of galena,5 sphalerite, and chalcocite. It is a well known fact that copper sul-phde minerals are oxidized slowly by weathering agents, so that the sulphides of copper which find their way into a tailing pond are gradually changed into water-soluble and acid-soluble compounds. Commercially, the oxidation of copper sulphide minerals is accomplished by roasting them at elevated temperatures in air using a suitable (usually multiple hearth) roaster. Most copper smelters are equipped with such roasters. In this phase of the study we were especially interested in determining the effect of temperature on: (I) the air oxidation of chalcocite and (2) the oxidation of chalcocite using commercial oxygen. The amount of sulphur removed as sulphur dioxide, the percentages of water-soluble copper, and of acid-soluble copper were determined for the air oxidation tests. In the tests using commercial oxygen, the percentages of sulphur dioxide, water-soluble copper, acid-soluble copper, water-soluble sulphur and acid-soluble sulphur were determined. For these experiments the temperature was varied from 250 to 700°C. Experimental Preparalion of Chalcocite (Cu2S) Natural chalcocite is not pure, Minerals such as bornite, covellite, and chalcopy-rite may be present. It was decided, therefore, to prepare some chalcocite in our own laboratory. An intimate mixture of sulphur and hool divided metallic copper in the ratio of 4 g of copper to I g of sulphur was prepared. TO this was added a slight excess of

Jan 1, 1949

By E. J. Schmitz

The mining of the Triassic coals of the Richmond basin has been carried on, upon a larger or smaller scale, for more than a hundred years. Notwithstanding the close proximity of the field to a large city and to tide-water, assuring both a good home-market and cheap transportation, and notwithstanding the good quality of the fuel and the great thickness of the principal bed developed and worked, operations in this field have been conducted with but limited success. The basin comprises about 189 square miles, being from 24 to 31½ miles in length by 5½ to 10 miles in width. Not much more than one-tenth of this area has been explored in a practical and efficient way, and not over 1200 acres have been worked near and along the outcrops or edges of the coal. The practical mistakes which have been committed undoubtedly in a number of the mine-operations of the past, do not alone explain the unsatisfactory success and progress of the industry of this field ; nor can it be ascribed to the lack of local enterprise. Other reasons have to be looked for; and, as a principal one, the irregularity and the faulted conditions of the coal deposits must be pointed out. As a second cause, the indifference of the State government, which leaves the development of the field entirely to private enterprise, instead of aiding it by systematical and practical geological investigations, deserves to be mentioned. Nothing has been done by the State since the excellent work done more than fifty years ago by Professor Rogers, who could not at that time enter into the details of such

Jan 1, 1895

By Ta M. Li

US mine-related fatalities were reduced by 10.3% at metal and nonmetal operations in 1974, according to a report by the Secretary of the Interior to Congress concerning the administration of Public Law 89-577. While there was no significant change in the overall injury experience, the 157 fatalities were the lowest recorded in the mining industry since federal mine inspections began under the Act in 1970. The report attributes this reduction to the "improvement in the quality and increased size of the inspection force as well as the continuing emphasis on mine safety conditions and practices." The year 1974 represents the first full calendar year in which the Mining Enforcement and Safety Administration (MESA) had responsibility for administration of the Federal Metal and Nonmetallic Mine Safety Act, and the Federal Coal Mine Health and Safety Act of 1969. Prior to MESA's establishment in 1973, these functions were delegated to the US Bureau of Mines. Research activities in health and safety were continued by the Bureau in 1974.

Jan 12, 1975

By J. G. Bradley

On February 1, 1971, United States Steel Corporation, Lorain Works began operation of its first BOP facility equipped with a partial combustion gas cleaning system.- This shop consists of two 220 ton vessels. Figure I schematically describes this partial combustion gas cleaning system and highlights the necessary features for it-s safe and efficient operation. Waste gases are drawn into the membrane hood by a 2250 H.P. induced draft fan and are first quenched by the six inch bull nozzle at the top of the hood. Larger solid particles are separated from the gas in the quencher venturi and final scrubbing takes place in the P-A high energy venturi scrubber before the cleaned waste gas is burned in the flare stack. Scrubbing water is routed through a cyclone separator and then cycled into one of two 100' diameter thickeners. Underflow sludge which averages 20% solids is dewatered by 100 H.P. centrifuges. The clear overflow from the thickeners is recycled first to the P-A venturi and then to the quencher before returning to the thickeners.

Jan 1, 1972

By R. J. Batterham, P. W. Roller

Granulated sinter mixes of three iron ore types were prepared with a range of water contents in a laboratory investigation. The mean diameter of the granulated mixes and the void fraction of beds of the mixes were determined experimentally. Pressure drop measurements over a range of flow rates were made in a laboratory permeameter with pressure tappings at several depths in the bed. The Ergun equation was applied to the results to determine the shape factor of the granulated mixes. Pressure readings down the bed were used to estimate the local void fraction in the bed using the Ergun equation and the calculated granule shape factors. It was found that there was a decrease in void fraction with depth in the bed. The significance of this and the variation of shape factor values with the amount of water used for granulation is examined in relation to the strength of granules of the sinter mixes. The experimental and mathematical techniques presented may be useful in interpreting the behavior of the permeability of sinter beds after ignition.

Jan 1, 1986

By J. C. Miller

In offshore oil drilling as it is done today, accidental blowouts cause considerable damage lo the environment. Public alarm over such accidents has already resulted in a number of legislative proposals which appear to favor the cessation of all drilling on the Federally owned continental shelf. Present concern for the environment and the increasing water depths associated with offshore drilling suggest the desirability of applying underground mining technology to offshore petroleum exploration. If offshore drilling were to proceed from underground accesses to the continental shelf, accidental damage to surface waters would be avoided, and so would the difficulties of exploring in deep or ice-filled waters. The specific approach proposed here is intended as a preliminary to offshore oil exploration in a known field or geologic structure. The end result of this method would be an on-shore vertical or inclined shaft terminating in an operating tunnel or haulageway as shown in the illustrations. From the haulageway, branching access tunnels would lead to drilling compartments. With this arrangement, it appears likely that one operating tunnel could serve all lessees ill a particular unit operation. Although an inclined shaft is not illustrated, this would probably be more practical than a vertical shaft owing to the length of the casings and the character of other equipment used in oil producing operations.

Jan 1, 1971

By C. W. Phillips, R. S. Busk

With most cast metals the grain size may vary within wide limits, depending upon the conditions at the moment of freezing. These conditions are subject to control in magnesium-base alloys, by proper melting and superheating techniques, enabling production of quite uniform and fine-grained castings, almost independent of section thickness. However, such variations of grain size as do exist, in common with all metals, are reflected in corresponding changes in mechanical properties. It is the intent of this paper to present data giving the relationship between grain size and mechanical properties. Data are also included on the combined effects of grain size and micro-porosity. In addition, a short discussion of factors influencing the grain size of sand castings is included. Measurement OF Grain Size By grain size is meant the statistical distribution of volumes occupied by the individual crystallites in a metal section. Since this is almost always impossible of direct measurement, several methods of estimating and expressing the size have been evolved. Most exact methods make use of the microscope. This involves study of a single plane cut at random through the metal. What is seen in the microscopc are polygons, each the cross section of a grain. The diameters or areas of these polygons are then measured and the grain size is expressed as an average diameter, number of grains per unit area, average area per grain, or, rarely, average volume per grain. Rutherford, Abom, and Bain3 have discussed extensively the errors involved when using a plane section to estimate volume. They clearly point out that the tendency is to underestimate the true grain size because of the nature of the measurcment. The A.S.T.M. recommends that measurement be made by comparison with a standard chart, by means of Jeffries' planimetric method4 or Heyn's intercept method. The Society recommends that the notation be in terms of the number of grains per unit area (as is the grain-size number used for steel) or as an average diameter, expressed in inches or millimeters (as for copper and brass). The method of measurement used for this study was comparison with a standard chart, which has been described by P. F. George.5 The chart was constructed by photographing at Ioo diameters a uniform specimen with an average grain diameter of 0.003 in. The average diameter of this specimen was determined by numerous measurements, using both the Jeffries and Heyn methods. The other grain sizes were then obtained by appropriate enlargemcnt of the master sample. Careful comparison of the chart method with other methods on identical samples indicates that though the differences be-

Jan 1, 1945