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By AIME AIME

THE Coal Division holds its fall meeting in the Pocahontas coal field, at the West Virginian Hotel, Bluefield, W. Va., Oct. 9 and 10. The first day will be a busy one-two sessions for the presentation of papers and a dinner in the evening. Hon. E. Fudge, Assistant Undersecretary, Mines Department, and R. V. Wheeler, head of the Mines Research Board of Great Britain, will be the guests of the Division. Mr. Wheeler will discuss the British Mining Act as relates to the coal industry. George S. Rice, Chief Mining Engineer, U. S. Bureau of Mines, will present an account of the recent meeting at Buxton in England, of representatives of England, Germany, France, Belgium and the United States to formulate a plan and to correlate research work now in progress in the various countries. The second day the members will be the guests of the operators on a trip through the Pocahontas field. The work of preparation for the meeting is being actively pushed by committees headed by T. H. Clagett, R. M. Lambie and H. D. Smith, and a good attendance, both local and from the general membership of the Division, is anticipated. Because many will make the trip by car the principal automobile as well as rail-road connections to Bluefield are given here. The program for Friday, Oct. 9, follows:

Jan 1, 1931

By GERALD A. WARING

ALASKA'S coal consumption is now about 130,000 tons annually. About one-quarter of this amount is used in the southeastern part of the territory and in settlements on the western coast and comes from Seattle and from British Columbia mines. About 60,000 tons is sub-bituminous coal from the Healy River mine in central Alaska, and is used mainly at Fairbanks, sixty miles to the north, for heating and the development of electric energy, for the city and the gold-dredging operations of the Fairbanks Exploration Co. The remaining 40,000 or 50,000 tons is bituminous coal from Matanuska Valley, in the south-central part of the territory. These coal deposits extend intermittently for about forty miles along the valley, which trends eastward from the upper end of Cook Inlet.

Jan 1, 1934

By K. D. McBean, W. H. Hannay, B. A. Stimmel

The electrolytic plant at Trail, B. C., was erected by the Consolidated Mining and Smelting Company of Canada, Ltd., for the treatment of zinc ores and concentrates, particularly those of the Sullivan mine at Kimberley, B. C. It began operations in 1916 at a rated capacity of 60 tons of cathodes per day, which has been steadily increased to the present production of 400 tons per day. The Sullivan has furnished the greater part of the zinc so far produced, although notable tonnages of customs zinc concentrates have been treated. The maximum yearly tonnage of customs concentrates, 25,000 tons, was treated in 1926. Practically no customs concentrates hale been treated since 1929. The total bar zinc production to the close of 1935 was 1,092,782 tons. The zinc mineral in the Sullivan ore is marmatite of somewhat variable composition, ranging between 6ZnS.FeS and 8 ZnS.FeS. Practically all the zinc ores tributary to this plant are marmatites; and the iron content of the zinc mineral varies considerably. The formation of zinc ferrite during roasting is now a matter of common knowledge, so the possible extraction from any zinc concentrate is largely a function of the iron content; the progress made in concentration is reflected directly in zinc-plant recovery. The Trail plant operafes on the low-acid system, as do all the other major plants on this continent at present operating, except the electrolytic plant of the Sullivan Mining Co. at Kellogg, Idaho. The new Magdeburg plant, in Germany, occupies an intermediate position between high and low acid. Plant operation at Trail will be treated in this paper under five main sections: roasting, leaching and purification, electrolysis, cadmium recovery and melting. The oxide section of the plant, treating zinc fume from re-treatment of lead blast-furnace slag, is practically a separate zinc plant; this is necessary because of the presence of fluorine in the smelter zinc fume. Cadmium is recovered electrolytically from the purification residues of both the sulphide and oxide leaching plants.

Jan 1, 1937

By M. A. Bryant, J. C. Goodwin, T. B. Faddies, L. D. Kornze

The Mercur Gold Mine is located 56 km (35 air miles) southwest of Salt Lake City in north central Utah; In July 1981 Getty Mining Company made the decision to put Mercur into production. The first gold bar was poured in April 1983. This paper provides details on: 1) the history of the camp, 2) past production, 3) geology of the deposit, 4) the exploration and development program, 5) ore reserve estimates, 6) detailed geologic pit mapping, 7) production blasthole sampling techniques and 8) the use of geostatistics (kriging) in handling assays and gold extractions to produce the required mill feed.

Jan 1, 1985

By John Greenawalt

IN view of the increasing importance of sintering in the beneficiation of iron ores preparatory to their reduction in the blast furnace, the writer believes the time is opportune for an up-to-date, though brief, discussion of the general principles of sintering, as well as a discussion of some of the recent improvements and mechanical developments that have been achieved in this industry, its present trend and future possibilities. Sintering may be defined as the agglomeration of fine mineral particles into a porous mass by incipient fusion caused by heat produced by com-bustion within the mass itself. Its first successful application, in connec-tion with the treatment of lead sulphide ores, was known as blast roasting. In this process the lead ore was given a preliminary roasting in a reverber-atory furnace, then the incandescent mass was placed in a pot with a grate and an updraft was applied. It was soon discovered that much better results were obtained by cooling the ore, moistening it, and placing it upon an incandescent layer of fuel previously placed upon a grate, and then applying the updraft blast of air until the charge was converted into a more or less agglomerated mass. The fuel was sulphur and the process one of desulphurization, which was the principal result desired. The application of downdraft to blast roasting brought about tremen-dous results; the principal one of which was that an air blast of high pressure could be passed through a layer of ore deposited on a grate, or on a grate covered by a layer of coarse porous material, without disturbing the particles of which the layer is composed. The air current is compelled to filter through the pores of the charge, whereas if the same pressure applied to the top surface of the charge with downdraft were applied to the bottom of the charge with updraft, the charge would be completely blown off the grate. Another result of reversing the draft was that a flame could readily be applied to the top surface of the charge for ignit-ing purposes.

Jan 1, 1938

By Geoffrey Gilbert

The Montana Phosphate Products Co., subsidiary of the Consolidated Mining and Smelting Company of Canada, operates three phosphate properties north and northeast of Garrison, Powell County, Mont. Production of all three is from the same phosphate bed, a 3 to 4-ft. member of the Phosphoria formation, whose outcrop is traceable across the mountains of the Garnet Range in a series of southeasterly pitching folds. Because of topographical differences and of variations in the attitude of the bed, each property has its own mining methods and problems. Most of the phosphate rock produced since 1930 has been shipped to the plant of the Consolidated Mining and Smelting Co. at Trail, B.C., where it is treated with sulphuric acid for conversion to phosphoric acid, which in turn is used to produce various fertilizers. History The occurrence of rock phosphate in Montana has been known since 1910, and the Garrison area was mapped by the United States Geological Survey in 1913. In the nineteen twenties William Anderson, owner of a ranch 8 miles north of Garrison, discovered phosphate on his land, and the Montana Phosphate Co., which later became the Montana Phosphate Products Co., was formed to work the deposit. Montana Phosphate Products Co. later obtained United States leases on two adjoining parcels of ground, and this combined property, known as the Anderson mine, has been the main source of Con-solidated's phosphate rock. The mine now includes an adjoining parcel leased from Anaconda and mined from the Anderson workings. In 1940, the Graveley mine, a small privately owned property 6 miles east of the Anderson, was leased from the owner by Montana Phosphate Products Co., and a U. S. lease held by Graveley on some adjoining ground was subleased. In 1943 the Luke mine, on the Mineral Hill property, 1 3/4 miles south-southwest of the Graveley, a U. S. lease, which previously had produced a few thousand tons, was optioned from the lessees by Montana Phosphate Products Co., and is now being developed. At both these mines the rock is appreciably higher grade than at the Anderson mine. Geology The Phosphoria formation is the Permian member of an extensive sequence of Paleozoic and Mesozoic rocks. It is underlain by Pennsylvanian (Quadrant) quartzite and Mississippian limestone, and overlain without any marked unconformity by Jurassic (Ellis) shale and Cretaceous shale and sandstone. In the Garrison area, the Phosphoria itself consists mainly of about 50 ft. of gray to yellow chert. Immediately below the chert is the phosphate, usually 3 to 4 ft. thick. Below the phosphate occur a few

Jan 1, 1948

By Geoffrey Gilbert

The Montana Phosphate Products Co., subsidiary of the Consolidated Mining and Smelting Company of Canada, operates three phosphate properties north and northeast of Garrison, Powell County, Mont. Production of all three is from the same phosphate bed, a 3 to 4-ft. member of the Phosphoria formation, whose outcrop is traceable across the mountains of the Garnet Range in a series of southeasterly pitching folds. Because of topographical differences and of variations in the attitude of the bed, each property has its own mining methods and problems. Most of the phosphate rock produced since 1930 has been shipped to the plant of the Consolidated Mining and Smelting Co. at Trail, B.C., where it is treated with sulphuric acid for conversion to phosphoric acid, which in turn is used to produce various fertilizers. History The occurrence of rock phosphate in Montana has been known since 1910, and the Garrison area was mapped by the United States Geological Survey in 1913. In the nineteen twenties William Anderson, owner of a ranch 8 miles north of Garrison, discovered phosphate on his land, and the Montana Phosphate Co., which later became the Montana Phosphate Products Co., was formed to work the deposit. Montana Phosphate Products Co. later obtained United States leases on two adjoining parcels of ground, and this combined property, known as the Anderson mine, has been the main source of Con-solidated's phosphate rock. The mine now includes an adjoining parcel leased from Anaconda and mined from the Anderson workings. In 1940, the Graveley mine, a small privately owned property 6 miles east of the Anderson, was leased from the owner by Montana Phosphate Products Co., and a U. S. lease held by Graveley on some adjoining ground was subleased. In 1943 the Luke mine, on the Mineral Hill property, 1 3/4 miles south-southwest of the Graveley, a U. S. lease, which previously had produced a few thousand tons, was optioned from the lessees by Montana Phosphate Products Co., and is now being developed. At both these mines the rock is appreciably higher grade than at the Anderson mine. Geology The Phosphoria formation is the Permian member of an extensive sequence of Paleozoic and Mesozoic rocks. It is underlain by Pennsylvanian (Quadrant) quartzite and Mississippian limestone, and overlain without any marked unconformity by Jurassic (Ellis) shale and Cretaceous shale and sandstone. In the Garrison area, the Phosphoria itself consists mainly of about 50 ft. of gray to yellow chert. Immediately below the chert is the phosphate, usually 3 to 4 ft. thick. Below the phosphate occur a few

Jan 1, 1948

By Michael D. Martin

Pima Mining Co., managed and 50% owned by Cyprus Mines Corp., with Union Oil of California and Utah Construction and Mining Co. sharing equally in the remaining 50%, is a medium-sized open pit copper mine situated in the very active Twin Buttes mining district, 22 miles southwest of Tucson, Ariz. In 1965, a three-part paper described Pima's growth from 3,000 tpd (tons per day) in 1957 to 18,000 tpd nine years later. Within months of the completion of the last stage of this sixfold expansion, studies were being wound up on a further major expansion, this time to 30,000 tpd. This expansion is now a reality. Details of the enlargement of the crushing and concentrating plant have already been covered in a recent paper, and the present paper will cover pit engineering studies, equipment selection, and current operating practices. Key factors in top management's decision to go ahead with the latest expansion were: (1) a substantial increase in ore reserves coupled with an appreciable decrease in grade, (2) a generally favorable outlook for copper prices, and (3) better than expected performance by mine and mill equipment purchased for the expansion to 18,000 tpd. For the pit design studies, particular attention was paid to pit stability, a problem area at times in the past, and to the location and grades of the final ramp systems. In drawing up the long-term mining schedule, the attitude of the ore, with the higher grade material lying deeper, was a critical factor and considerable stockpiling of lower-grade material in the early years had to be accepted in order to keep the annual grades close to the average of the orebody. The expansion called for major purchases of mine equipment in order to gear up the mine for delivery of 30,000 tpd to the concentrator. In general, equipment chosen was of types already proven at Pima, al- though there were some exceptions to this practice. One of these excep-

Jan 1, 1969

By B. W. Howlett, M. B. Bever, Somnath Misra

The heats oj- formation at 0°C of the compounds AuSn,, AuSn2, AuSn, and the £ phase were measured in a metal -solution calorimeter with liquid tin or a liquid Sn-Bi alloy as solvent. The melting point, heat of fusion, and heat capacities near the melting point of- the compound AuSn in the solid and liquid state were measured in a constant temperature gradient calorimeter. The free energy and entropy of formation of the compound AuSn at its melting point and at 298 °K were calculated from the experimental results and published data. The heats of formation are discussed in relation to the nature of the compounds. The heat capacity below the melting point, heat of fusion, and entropy of fusion of AuSn are interpreted in terms of processes oc-curring before melting. The atomic arrangement in liquid Au-Sn alloys and the bonding in the intermediate phases of. the Au-Sn system are discussed in the light of the measured thermodynamic properties. THE available thermodynamic information on the intermediate phases in the Au-Sn system consists of the heats of formation of the compounds AuSn4, AuSn2, and AuSnly2 and of one composition of the < phase2 besides values of the heat content of the compound AuSn from 0" to 635".&apos; However, the heats of formation refer to different temperatures ringing from 90" to 450°C. Also, the reported error limits of some of the values are large and several inconsistencies are apparent. In the investigation reported here the heats of formation of the compounds AuS-, AuSnz, and AuSn and of three compositions of the < phase were determined for a single reference temperature of 0°C by liquid metal-solution calorimetry. The compound AuSn, which is the only congruently melting intermediate phase in this system, was further investigated in a constant temperature gradient calorimeter by measuring its heat of fusion, melting point, and heat capacity in the solid and liquid states near the melting point. From these data and published free energies of the liquid of composition Au~n,&apos; the free energies of formation of the compound AuSn at the melting point and at 298°K were calculated. The results shed light on the solid-liquid transition of AuSn and on the atomic bonding and state of association of gold and tin in the solid and liquid phases. 1) EXPERIMENTAL PROCEDURES 1.1) Preparation of Samples. Stoichiometric amounts of gold and tin, each 99.99+ pct pure, were melted in evacuated Vycor capsules (12 mm ID). The melts were maintained approximately 150°C above the liquidus for about 12 hr and shaken repeatedly. Samples of the compound AuSn were cooled slowly to room temperature. Melts of the other phases, all of which solidify by peritectic reactions, were transferred to the narrow end (5 mm ID) of the capsule, which was then quenched into iced brine. The solidified samples were annealed in the unopened capsules for 10 days at 30" to 50° C below the respective solidus temperatures. The absence of microsegregation and second phases was confirmed by metallographic examination. At least two batches of each composition were prepared. 1.2) Solution Calorimetry. The heats of formation were determined by liquid metal-solution calorimetry. In this technique, the difference between the heat effects on dissolution in a liquid metal, or alloy bath, of the compound and a mechanical mixture of its annealed components, adjusted for changes in composition of the bath, is the heat of formation of the compound at the temperature from which the samples are added. Details of the equipment and technique have been described."&apos; A tin bath at 300" or 350°C was used with the compounds AuSn,, AuSnz, and AuSn. The high gold content of the £ phase would have caused it to dissolve with an appreciable exothermic heat effect. In order to achieve thermal compensation, the samples were added from 0°C to a bath at 350°C consisting of 66.2 wt pct Bi (99.999+ pct pure), balance Sn (99.99+ pct pure). This solvent gives thermal compensation because gold dissolves exothermically in tin and endothermically in bismuth. The calorimeter was calibrated by adding from 0°C approximately 0.015 g-atom of the solvent.

Jan 1, 1965

By Harold E. Rowen

Sintering is accomplished at a temperature of more than 2000°F. For the purpose of this discussion it will be defined as the art of burning a solid fuel with 90 to 95 pct ash content. Think of the problems of combustion involved in keeping a home furnace burning properly with 12 to 20 pct ash in the fuel and remember that the amount of fuel involved in the 5 to 10 pct combustion loss represented by the definition has to be just about right or nothing is going to happen. Our forefathers discovered that many ores containing metals csscntial to the steadily rising standards of living were associated with other minerals, sulfur and arsenic, for example, which had to be removed before ultimate refining. They did this by utilizing the solid fuels they had available-the sulfur or other readily combustible components- to create roasting heat.

Apr 1, 1956

By Rudolf Gahl

THE history of flotation in America is very short, at least as far as the large-scale application of the process is concerned. It is remarkable how many important developments have taken place inn the last few years and are already being extensively utilized. What was new a year ago, seems almost commonplace now. For this reason it is with hesitation that I follow the suggestion: of Dr. Ricketts, President of the Institute, to describe the experiences of the Inspiration Consolidated Copper Co. with the flotation process. TESTS CONDUCTED IN SMALL TEST-MILL When the Inspiration company first decided to build a concentrating plant, nothing was known about flotation, and the process was to be gravity concentration pure and simple. Demonstration Tests Conducted by Minerals Separation Co. While plans were being prepared by H. Kenyon Burch (who had been intrusted with the design and construction of the concentrator) the Minerals Separation Co., a concern at that time little known in America, asked and obtained permission to demonstrate the value of its flotation process for Inspiration ore. As. a consequence, a small 50-ton flotation machine of standard design was added to the company&apos;s test plant and started to operate in the beginning of 1913. This marks the beginning of flotation at Inspiration. The results obtained with this machine, which was operated by members of the Minerals Separation staff, so far surpassed what this company anticipated that it was decided to continue flotation tests for this purpose, and two of the flotation experts of the Minerals Separation Co., I. L. Greninger and G. A. Chapman, were retained. L. R. Wallace, now superintendent of the Miami works of the International Smelting Co., was at that time metallurgist of the Inspiration company and in that capacity took an active part&apos; in these tests. Great credit is due to him for his quick recognition of the possibilities of flotation.

Jan 9, 1916

By Gerhard Derge

A series of studies of the corrosion of tin is under way in the Metals Research Laboratory at the Carnegie Institute of Technology. The complete program includes examination of the corrosion properties of high-purity tin and the changes produced in these properties by the addition of varying amounts of alloying elements. Other studies of this subject have been made recently under the sponsorship of the International Tin Research and Development Council, but these researches have been confined largely to the weakly acid environments commonly encountered in the canning industry.&apos; There is little information available regarding the corrosion of tin under alkaline conditions, but such data would be of value, for a large amount of the metal is used under such conditions. The potential measurements to be described were made because they seemed to offer the most rapid method for obtaining a general view of the conditions to be encountered in such service. These experiments will also serve to establish a technique that will be of value in further studies. It must be emphasized that the results reported are those of controlled laboratory experiments and that the conclusions drawn cannot yet be applied to service conditions; rather, they point to the directions in which more exhaustive experimental work may be done to best advantage. Experimental Procedure The first work was devoted entirely to the problem of obtaining reproducible results from a series of potential measurements. Only the methods finally adopted as standard will be described here, except when the variations from these have furnished results of interest. The cell used is shown schematically in Fig. 1. The electrolyte is contained in the vessel A (an inverted bell jar), whose ground top is fitted to the transite-board lid B. The controlled atmosphere can be admitted

Jan 1, 1938

By Gerhard Derge

A series of studies of the corrosion of tin is under way in the Metals Research Laboratory at the Carnegie Institute of Technology. The complete program includes examination of the corrosion properties of high-purity tin and the changes produced in these properties by the addition of varying amounts of alloying elements. Other studies of this subject have been made recently under the sponsorship of the International Tin Research and Development Council, but these researches have been confined largely to the weakly acid environments commonly encountered in the canning industry.&apos; There is little information available regarding the corrosion of tin under alkaline conditions, but such data would be of value, for a large amount of the metal is used under such conditions. The potential measurements to be described were made because they seemed to offer the most rapid method for obtaining a general view of the conditions to be encountered in such service. These experiments will also serve to establish a technique that will be of value in further studies. It must be emphasized that the results reported are those of controlled laboratory experiments and that the conclusions drawn cannot yet be applied to service conditions; rather, they point to the directions in which more exhaustive experimental work may be done to best advantage. Experimental Procedure The first work was devoted entirely to the problem of obtaining reproducible results from a series of potential measurements. Only the methods finally adopted as standard will be described here, except when the variations from these have furnished results of interest. The cell used is shown schematically in Fig. 1. The electrolyte is contained in the vessel A (an inverted bell jar), whose ground top is fitted to the transite-board lid B. The controlled atmosphere can be admitted

Jan 1, 1938

By Hansheinz Portisch, Harold Margolin

A surface expansion was found to occur sometime after etching in Ti-A1 alloys containing 9.5 to 12.5 wt pct Al. The structure formed, grew, and disappeared with tzrrze. The surface expansion was followed by microscope observations and interferometric and lattice parameter measurements. Activation energy measurements for the growth of the "expansion structure" and chemical analysis indicated that the phenomenon occurred as a result of hydrogen pzckup during etching. It is proposed that hydrogen initially enters octahedral sites of Ti3Al coherent with a Ti and later shifts to the tetrahedral sites. It is postulated that expansion occurs when hydrogen enters the tetrahedral sites. The expansion structure disappeared, it is proposed, because of diffusion of hydrogen from the surface into the body of the alloy and because of loss of coherency of Ti3Al. In examining Ti-A1 alloys, Ence and arolinl observed markings on the surface of specimens. These markings did not appear after electropolishing only, but rather appeared only after etching. The markings appeared to grow as a function of time after etching and later seemed to disappear. Although the markings had some similarity to precipitates, showing, frequently, a Widmanstatten type of arrangement, the observation that other microstructural markings continue to be seen within the new structure suggested that it was actually not a precipitate. The source of this structure was unknown and an attempt was made in the present investigation to develop some understanding of its nature. It has been labeled expansion structure. 1) EXPERIMENTAL PROCEDURE A) Alloys. Alloys in the range 6.5 to 14.5 wt. pct A1 were studied. Bars, 4 in. sq, forged from Bureau of Mines (73 Bhn) titanium consumably arc-melted 4-lb ingots,&apos; as well as 50-g arc-melted buttons of desired aluminum contents were used. The 50-g buttons also used Bureau of Mines titanium (73 Bhn) and aluminum of 99.99 pct purity. Buttons containing up to 8.5 wt pct A1 were hot-rolled from a furnace at 900&apos; . Those with higher aluminum contents were hand-forged on a titanium anvil, and heated with an oxygen-hydrogen torch in the region of 1200" to 1300°c. Frequent reheating kept the samples at the desired temperature range. After a reduction of about 30 pct, the samples were water-quenched. To eliminate any contamination picked up either during hot rolling or forging, at least 1 mil of the surface of the sample was taken off. B) Heat Treatment. Prior to heat treatment all alloys were vacuum-annealed to remove hydrogen. Samples were annealed at 900&apos; until a vacuum of 10&apos;5 mm Hg was established. After this treatment, the samples were wrapped in molybdenum sheet and heat-treated in argon-filled quartz capsules, which were broken under water or iced brine at the conclusion of the heat treatment. All heat treatments under dynamic vacuum were performed in a rapid-quench furnace. This consisted of a molybdenum-lined quartz tube attached to a vacuum system and through a stopcock to a beaker of water. At the completion of the heat treatment the vacuum stopcock was closed, the furnace shut off, and dropped below the quartz tube. Then immediately the inlet stopcock was opened and water admitted until the tube was filled. The steam formed was allowed to escape through the inlet stopcock. This method was used in heat treatments up to llOO°C. C) Metallography. Specimens for metallographic examination were ground, then electropolished using a Disa Electropol machine with a perchloric acid electr01te. Specimens were etched with R-etch.3 A standard etching time of 3 min was used, with the specimen being agitated during immersion. D) Sample Preparation for X-ray Analysis. Samples 0.2 to 0.5 mm in diam were produced from heat-treated rods which were turned to 2 mm diam on a lathe and then rotation-etched. An etchant consisting of 1 pt HF, 7 pt HN03, and 12 pt HzO was satisfactory. With the rod rotating in a vertical position at 50 to 100 rpm, a needle with uniform dimensions could be obtained. Care had to be taken to insure that the rod was at the center of rotation, otherwise cavitation developed. If a small unevenness developed, it was possible to grind it off with a fine emery paper. E) X-ray Diffraction. All X-ray work was done on a North American Phillips X-ray diffraction apparatus and a Jarrell Ash microfocus X-ray unit. The Phillips unit was used with a copper target and nickel filter. The Jarrell Ash unit was fitted with a cobalt target and iron filter. For the Phillips unit 35 kva and 20 ma were used, whereas for the microfocus unit with the 100-p fixed-focus gun 40 kv and 1.5 ma were used. It was found that alignment of cameras on the Jarre11 Ash unit was very critical. The X-ray beam contains an intense area which is not the beam center. The cameras were aligned with the intense region by monitoring the beam coming out of the camera with a Geiger counter. Adjustments were made until a maximum intensity was obtained. A Phillips diffractometer with a Brown chart recorder furnished some of the lattice parameter data. The divergence slit up to 80 deg 28 was I deg; above 80 deg, a 4-deg opening was used, while the scatter slit was1 deg and the receiving slit had a 0.003-in. opening. The general scanning rate was 1 deg per min, while peaks of special interest were rescanned at -j deg per min. For elevated-temperature X-ray diffraction a Uni-cam High Temperature Camera with a film diameter

Jan 1, 1969

By L. H. E. Weyher

As a result of past developments, mainly at universities, the coal industry has had access for a decade or more to a number of computer programs for coal mine planning. Using some of these programs Bethlehem Steel&apos;s Research Department in 1969 launched jointly with Bethlehem Mines Corporation an extensive research and development program aimed at: (1) testing the capabilities of these programs through trial applications as the need and opportunity existed, (2) defining what additional capabilities were required to make mine planning computer systems fully useful in the hands of Bethlehem&apos;s mine-planning engineers in the field, and (3) developing and implementing in the field the mine-planning systems agreed on. This program is now nearing completion, and several of Bethlehem&apos;s operations routinely use the planning tools that were developed. The progress of this research program is discussed in terms of our initial trial applications of computer methods to mine planning and the incorporation of several innovative features into our new set of computer programs, in particular, the use of time sharing, data base management systems, and general-purpose simulation languages.

Jan 1, 1977

James DouGlas, New York, N. Y. (communication to the Secretary*):—In my paper on the Conservation of Natural Resources, I referred to the slow replacement of bee-hive ovens by the by-product ovens as a most notable instance of waste. And I quoted from Mr. Parker&apos;s report for 1906 an explanation given by him in accounting for the small production of by-product coke. It was that the market for the by-products of the coke-ovens was so limited that some of the ovens constructed were out of operation. His report on the manufacture of coke in 1908&apos; does not record an improvement, and attributes the strange fact that we alone, of all the industrial peoples, delay the adoption of this cardinal improvement from the continuance of the same almost inexplicable cause. To quote again from his report, he says (p. 241): "The year 1908 was not marked by any notable gain in the construction of byproduct coking plants, though some new work was done. There was a net increase

Jan 1, 1910

By S. S. Solomons, V. H. Tillmann, L. Smaniotto

INTRODUCTION The Zinc Corporation, Limited and New Broken Hill Consolidated Limited (ZC/NBHC) mines are situated at Broken Hill, New South Wales, Australia (refer Fig. 1). Both companies are wholly owned subsidiaries of Australian Mining and Smelting Limited (AM&S). Annual production from both mines is 2.3 million tonnes of lead, silver and zinc ore mined by a variety of underground methods. From 1911 to 1968, conventional overhand cut and fill stoping (both open and square set timber) was used almost exclusively, though some trials with shrinkage and sublevel stoping were undertaken. The trial stopes used short holes in conparison with today&apos;s sublevel open stopes. Mechanised cut and fill methods with regular rockbolt support were introduced in the early 1960&apos;s and undercut and fill stoping replaced square set timber stoping for pillar extraction in the late 1960&apos;s. Since 1968, increasing emphasis has been placed on longhole open stoping (LHOS) although some areas of the orebodies are comrnrnitted to the abovemehtioned methods since existing stope development makes it impracticable to change. In 1979, approximately half the total production was mined by LHDS and it is expected that the LHOS proportion will increase in the future. A vertical crater retreat method of longhole mining has also been successfully introduced to replace the undercut and fill stoping of some pillars and it is expected that this stoping method will be used in some future primary ore extraction. This paper reviews the progress of LHOS technology at ZC/NBHC with emphasis on the stope planning and design process and the importance of rock mechanics applications. The objective has been to maximise the primary extraction of the orebody giving due consideration to stability and safety. This has led to the design of a novel sloped wall longhole stoping method which should eliminate the formation of pillars by mining the orebody in a "continuous" manner incorporating a sloped sand interface between the blocks of ore being mined and the previously mined sandfilled stope. The expected results include the complete extract of the orebody without the formation of pillars and with minimal dilution of ore with sandfill. The evolution of the method can be best illustrated by following the mining progress of the relatively new Southern &apos;B1 Lode orebody at the NBHC mine. GEOLOGY Southern &apos;8&apos; Lode is the southern extension of the &apos;Bt Lode orebody, which is currently being mined further to the north in ZC/NBHC (refer Fig. 2). The ore block being developed and stoped extends for a vertical height of 90 metres from No. 17 level down to approximately ten metres above No. 19 level. The ore minerals are sphalerite and galena with the zinc:lead ratio averaging approximately 3:l with a minor amount of silver associated with the galena. There is a sharp contact between the ore and the country rock on the footwall and keel, while the hanging wall gradates into medium and low grade material. The gangue is primarily quartz together with garnet, biotite, feldspar, gahnite, sericite and minor apatite. In various horizons within the orebody, grey feldspar, quartz and garnet pegmatites occur as stratiform layers. The ore and country rock are competent with some inconpetent boundaries.

Jan 1, 1981

By Richard McCaffery

THERE has been some discussion among blast-furnace operators regarding the relative merits of tuyeres of the converging type and tuyeres designed in an attempt to produce a diverging jet of air. An article1 on Progress in German Blast Furnace Practice indicates the interest in this subject abroad. This article illustrates a diverging type of tuyere which is shown here in Fig. 1A. The design appears to be rather poor as regards known properties of air flow. The adoption of this. tuyere, the article states, resulted in a decrease of blast pressure of about 4 lb. per sq. in. and an increase of the output of the furnace. If the improvements mentioned are due entirely to the change in tuyere design, the subject of tuyere design seemed worthy of investigation and this paper is a report of such work. Because of the gain in efficiency brought about by the use of a diverging tuyere in the German plant, it was thought desirable to investigate the shape of the air streams issuing from various-types of tuyeres to ascertain whether the gain was due to the tuyere itself or to the distribution of the air in the furnace. Since it is difficult to do experimental work on a full-sized furnace, or even on full-sized tuyeres, and because of the equipment available for the work, it was decided to carry out the work on scale models of tuyeres. Therefore, in this work, single tuyeres were used and the blast discharged into the atmosphere. The shape of the air stream was determined by means of a Pitot tube suitably supported in the air stream at various positions. The air was. furnished by a Roots blower driven by a 10-hp. direct-current motor having field control, and capable of discharging 250 cu. ft. per min. with a pressure on the discharge of 4 in. of mercury. A modern blast furnace with 21-ft. hearth requires about 50,000 cu. ft. per min. of free air at from 16 to 23 lb. per sq. in. pressure, depending upon the type of burden. This pressure is the total pressure required to force the air through the burden and through the tuyeres. Upon computation the pressure drop through the tuyeres is about 1 to 2 lb. per sq. in. By making the tuyeres to a one-fifth scale, and by using a pressure of 1.5 in. of mercury, the shape of the air streams should be

Jan 1, 1931

By D. W. Levinson, L. F. Mondolfo, N. A. Gjostein

MANY investigations are reported in the literature on the age hardening of Al-Mg-Zn alloys but most of them are concerned mainly with mechanical property changes. The present investigation was started to determine the structural changes in an AlMgZn alloy. The alloy composition was chosen in the existing phase diagram" so as to lie on the quasi-binary section Al-(AlZn)Mg Actually, it was found that the diagram as published was inexact at the lower temperatures and that the phase in equilibrium with aluminum at the lower temperatures is not the ternary compound. Technique The alloy used for the investigation was prepared from high purity materials and the resulting composition as determined by chemical and spectro-graphic analysis was: Zn, 6.17 pct; Mg, 2.13 pct; CU, 0.03 pct; Si, 0.009 pct; Fe, 0.005 pct; and the remainder, aluminum. An ingot -in. thick was cast, homogenized, hot and cold-rolled to strip 0.064-in. thick for dilato-metric and powder work, and to foil 0.005-in. thick for single crystal work. Solution treatment temperature for all specimens was 475°±5°C for periods of at least 1 hr, either in air furnaces or lead baths. Aging was done in oil or Pb-Bi baths controlled to within ±3°C. The aging was investigated at the following temperatures: room (30°), 130°, ZOO0, 260°, and 300°C. For the dilatometric study, the length changes were measured continuously, at temperature, using strips of alloy 10 in. long, shaped as channels to eliminate undesirable deflection due to bending. To compensate for thermal expansion and contraction due to the variations in the temperature of the aging baths, the dilatometers were built of the same alloy as the one being tested, with the same shape and size, but with material fully annealed. Gages on which 0.0005 in. could be read and 0.0001 in. estimated were used to measure the dilation. Length changes were also measured at room temperature. Strips approximately 10 in. long were solution-treated, quenched, and measured in a supermicrome-ter in which 0.0001 in. could be read and 0.00001 in. estimated. The specimens were immediately aged for a given time, remeasured, and then the cycle was repeated, including the solution treatment. The error of measurement with this technique was somewhat larger, but the results confirmed those obtained by continuous measurements. The sheet used for the dilatometric work was checked to determine grain orientation. Both X-rays and etching methods revealed a rather coarse grain size but only a negligible amount of preferred orientation. Thus, volume changes calculated from length changes are not invalidated by preferential orientation effects. Pieces of the sheet used for the dilatometric work were also used for the metal -lographic work. Filings for the powder X-ray study were produced from a homogenized piece of sheet. These filings were magnetically purged of iron, screened to 200 mesh, sealed in a Pyrex tube containing an argon atmosphere, and solution-treated. When quenching, the glass tube was broken, so that the powder dropped directly into cold water. After drying with alcohol, the powder was sealed into capillary tubes containing an argon atmosphere and then aged for the required times. In spite of all these precautions the three strongest lines of a( Al2o) appeared in all patterns, and were used as internal standards. By the use of these standards, accurate corrections for film shrinkage were obtained. For the back reflection patterns, annealed nickel powder was mixed with the specimen to act as internal standard. A preliminary investigation was made in small (57.3 mm diam) powder cameras. From these patterns the changes which took place in the solid-

Jan 1, 1957

By Luther Wagoner

Having given in a former paper1 the results of assays of sea-water, bay-mud, dredgings from San Francisco bay, etc., and believing it might be interesting to extend the work to include some deep-sea dredgings, I procured from the Smithsonian Institution six samples taken by the U. S. steamer Albatross, and marked as follows: Depth in No. station. Locality. Latitude. Longitude. Fathoms. 1. 2103. Off Delaware bay. 38&apos; 47&apos; 20" N., 72o 37&apos; W., 1,091. 2. 2265 to Between Chesapeake 35° 37&apos; N., 74O to 75O W., 49 to 70. 2297. bay and Hatteras. 3. 2420. Off Chesapeake bay. 37&apos; 03&apos; 20" N., 74O 31&apos; 40" W., 104. 4. 2528. East of Georges 41° 47&apos; N., 65o 37&apos; 30" W., 677. bank. 5. 2572. SE. of Georges bank. 40° 29&apos; N., 66O 041 W., 1,769. 6. 2681 South of Nantucket. 39&apos; 43&apos; N., 70&apos; 29&apos; W., 990. The above samples were assayed by the cyanide-method, described in my former paper (p. 807), about 30 g. of sample being used, and the following results were obtained: Value in Milligrams Per Metric Ton. No. Station. Gold. Silver. 1. 2103. 145 1,014 2. 2265 to 2297. 44 304 3. 2420. 15 353 4. 2628. 267 1,963 5. 2572. 125 377 6. 2681. 66 414 No. 4, Station 2528, consisting of red clay and volcanic ash, mas re-assayed ; 35 g. was roasted at a low heat for one hour and weighed 34.05 g. after roasting; the color was changed to a deep brick-red; and when assayed by the cyanide-method

Jan 1, 1908