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By G. R. Peterson, K. E. Porter

The U.S. Bureau of Mines has estimated the potential availability of copper from 204 mines and deposits in market economy countries (MEC's). The evaluated properties have demonstrated resources totaling 436.4 million metric tons of contained copper and account for 90% of the Bureau of Mimes reserve base for copper in market economy countries. Total recoverable MEC copper resources are 340.8 million metric tons, 69% of which is from producing mines and 31% from nonproducing mines and deposits. Chile had the lowest estimated average total cost from producing mines of W.48 per pound of recoverable copper at a 0% discounted cash-flow rate of return (DCFROR), with estimated average total costs ranging from $0.40 to $0.81 per pound. The estimated average total cost of production, per pound of copper, for producing mines in the United States amounts to $0.57 in January 1988 dollars at a Wo DCFROR, with estimated total costs ranging from $0.36 to $0.85 per pound. In both real and nominal terms, the United States has, on average, significantly lowered its copper production costs since 1981. Rationalization of the industry and significant increases in productivity have made a strong improvement in the competitiveness of the U.S. copper industry to the extent that the United States should no longer be considered as a marginal producer of copper.

Jan 1, 1992

By John F. Blandy

Every new mining district has had its own peculiar experiences in inventing and experimenting upon new methods for the various operations of mining, and more particularly in the processes of crushing and dressing ores. As a matter of course, during this period many old things have been reinvented, patented, and cast aside, there to remain until at a future day other geniuses shall bring them forth again as new. In this respect the Copper Region of Michigan has not been behind other mining centres, and probably in no other part of the country has more money been expended in devising new machines and improvements upon old ones for the crushing of the rock. The appliances for washing the sand have not been so varied, simply for the reason that, having but one mineral, or rather metal, of high specific gravity to separate from rock material which varies but little in its character in any one mine, it requires much less care than is necessary in most mining centres of the world. At the beginning of operations in that district most of the work was in the control of Cornish miners, who introduced the simplest of Cornish mills, namely, wooden stem pestles, with wooden shafts and cams. These were well suited to the small mines, and particularly to the limited means of transporting more expensive machinery. As these facilities improved, they were enabled to change to iron, and to vary their patterns of rods, shafts, cams, and mortars. The most approved pattern, finally obtained, is the square or round stem, with collar adjustable by means of keys. In its present form it is a bar of cold rolled shafting, an eye in the top, an adjustable collar with key-plate and keys, the head and a shoe of chilled iron. The stem is fitted into the head by a slight taper. The battery has heretofore been of wood, lined with chilled cast plates, and bed-plate of the same, but within the past year the California pattern has been introduced. The screens are of sheet-steel, drilled with 16 holes to the inch. During this period of trial and gradual improvement in the pestle stamps, in the years 1855-6, the Ball steam-stamps were introduced, and after years of labor and expense, have been made the most efficient and powerful machines ever yet used for the purpose. It is ostensibly the Nasmyth steam-hammer, and yet the

By S. Ramachandran, R. A. Walsh, J. C. Fulton

A method has been developed for calculating the oxygen content of commercially melted stainless steels. Application of thermodynamics to the de-oxidation reaction in straight chromium and Cr-Ni stainless steels has resulted in a set of equations DURING the past 30 years, many rewarding developments have resulted from the application of physical chemistry to steelmaking. A need still ex- which permits the calculation of oxygen contents. The model was programmed for computer calc~ilations. The validity of the method was demonstrated on routine oxygen samples from numerous commercially melted heats of stainless steel. ists for further progress in the use of thermodynamics to meet the more stringent requirements for cleaner steel. Current electric furnace practices have been developed around minimizing residual-oxygen content, one of the most important factors controlling the inclusion content of stainless steels. The residual oxygen is controlled by the use of deoxidizers during the finishing period and a number of empirical rules have been evolved. Although this feature of electric furnace practice has long been in use, only in recent years has funda-

Jan 1, 1963

By J. B. Clark, A. J. McEvily, R. L. Snyder

The nonuniform distribution of precipitate particles has been recognized as a leading factor contributing to the relatively low fatigue resistance of aluminum alloys. The structure of many of these alloys is characterized by narrow precipitate-free zones adjacent to the grain boundaries. Alloys with such zones exhibit a tendency for brittle inter crystalline fracture. The interrelation between this type of structure and mechanical properties was investigated in an Al-10 wt pct Mg alloy. It was found that deformation during fatigue occurs preferentially along these zones and cracks initiate there. In Al-10wt pct Mg, the zones were found to be supersaturated even after extensive general precipitation and are due to the absence of proper precipitate nuclei in the region near the grain boundaries. Cold working the alloy prior to aging improves the mechanical properties by inducing precipitation within the zones and also by jogging of grain boundaries. The mode of fracture is changed from brittle inter crystalline to more ductile trans granular fracture. THE process of fatigue is highly structure sensitive, with the strength of the whole often dependent upon some localized discontinuity, either geometrical or metallurgical in nature. Much has been learned about the role of geometrical discontinuities, e.g., notches, in fatigue, but with the exception of the effects of inclusions or the shapes of carbides, relatively little is known about the specific effects of discontinuities in metallurgical structure such as nonuniform precipitation. In most age-hardening aluminum alloys, metallo-graphic studies have shown that the extent of precipitation adjacent to grain boundaries is much less than that which occurs in the interior of the grains. The width of these almost precipitate-free regions, which are sometimes called denuded zones, and the extent of solute depletion within them, are dependent upon the particular alloy and its aging treatment. It has been observed1 that these zones are relatively soft with the result that plastic deformation takes place preferentially within them. It has also been shown 2-4 that there exists a tendency for intercrys- talline cracking in fatigue when such zones are present. It is of interest to note that Broom et al.2,3 were able to reduce the incidence of this type of failure in an A1-4 wt pct Cu alloy by stretching the material 10 pct prior to aging. In the present study, the effects of precipitate-free regions on the fatigue properties of an A1-10 wt pct Mg alloy were studied in detail, and the effects of deformation prior to aging on the nature of the precipitation process as well as on fatigue properties were also investigated. MATERIAL AND PROCESSING An A1-10 wt pct Mg alloy was selected for this study, because it was known that well-defined precipitate-free regions along the grain boundaries are readily obtained in this alloy after aging at 200oC.5 The starting materials were 99.998 pct A1 and singly sublimed magnesium of about 99.9 pct purity. The aluminum was induction melted in a graphite crucible, and then the magnesium addition was immersed until dissolved. Chlorine gas was then bubbled through the molten alloy for 4 min to degas the melt, after which the melt was cast at a pouring temperature of 730" to 760°C into a cold, graphite-coated, tapered steel mold. Since A1-Mg alloys are difficult to homogenize,5 special care was taken to obtain a uniform composition. Two-in. cubes were cut from the ingot and heated at 446°C for 30 min. These cubes were then hot forged approximately 35 pct in each of the three cube directions and homogenized for 16 hr at 446°C. Sheet specimens were then obtained by pressing 40 pct and rolling 35 pct per pass with reheating between reduction steps to a final thickness of approximately 0.10 in. The sheet was then solution treated for 16 hr at 446°C and water quenched. The age hardening behavior of this material at 200°C was then determined, and the results are shown in Fig. 1. The age hardening of this alloy when subjected to cold work prior to aging is also shown in this figure. Preliminary work indicated that extensive deformation after quenching was required to affect drastically the precipitate-free regions in this alloy, and a rolling reduction of 50 pct was chosen. For purposes of comparison the following three conditions were studied: a) Solution treated, quenched, and aged 20 hr at 200°C

Jan 1, 1963

By R. E. Whitehead

"Abstract - Biotite from aplite that appears to have crystallized simultaneously with disseminated Cu-Mo mineralization at MacLeod Lake in northern Quebec yields K-Ar dates in the range 2032 ± 44 Ma to 2097 ± 47 Ma. The mean value, 2065 Ma, is some 400 Ma younger than the 2485 ± 64 Ma date obtained from a hornblende concentrate separated from a granodiorite sill that is spatially, but apparently not genetically, associated with the disseminated Cu-Mo deposit. The mean biotite date of 2065 Ma at MacLeod Lake, a K-Ar date of 2055 Ma reported from an area 50 km east of MacLeod Lake and published dates of about 2000 Ma for galenas from the Chibougamau Cu-Au ores, suggest a widespread thermal and mineralization event, in that part of the Superior Province east of James Bay, that differs from the common late Archean mineralization that is related to volcanism and metamorphic-intrusive activity.Sommaire - Les datations au K-Ar sur des biotites qui auraient cristallise en même temps que la minéralisation disséminée en Cu et Mo dans une aplite mise en place au Lac MacLeod, au nord du Québec, s'échelonnent entre 2032 ± 44 Ma et 2097 ± 47 Ma. La datation moyenne, soit 2065 Ma, est d'environ 400 Ma plus récente que la datation obtenue pour un concentre de hornblende extrait d'un filon-couche de granodiorite qui est spatialement mais ne serait pas génétiquement associée à la minéralisation disséminée en Cu et Mo. La datation moyenne de la minéralisation au Lac MacLeod, soit 2065 Ma, la dataion au K-Ar à 2055 Ma citée pour une région à 50 km à l'est du Lac MacLeod, de même que les datations d'environ 2000 Ma établies pour des galènes qui se trouvent dans le minerai aurifère et cuprifère à Chibougamau, suggèrent un phénomène thermique répandu qui aurait produit la minéralisation dans cette région à l'est de la Baie James, dans la province du lac Supérieur. Ce phénomène serait diffèrent de celui qui aurait produit la minéralisation répandue d’âge archéen tardif qui elle serait associée à l’activité volcanique et à l‘activité intrusive accompagnée de métamorphisme."

Jan 1, 1992

By Philomene Verlaan

The Code for Environmental Management of Marine Mining (the Code) was adopted in 2001 by the International Marine Minerals Society (IMMS) following extensive and wide-ranging consultation, and updated in 2002. The Code provides a framework and benchmarks for development and implementation of an environmental program for a marine exploration or extraction site by marine mining companies, and for stakeholders in governments, NGOs and communities in evaluating actual and proposed applications of environmental programs at marine mining sites. The Code provides for its review by IMMS after five years, with a focus toward possible modifications based on experience with its implementation and in light of developments in the field of marine mining and associated environmental practices, in consultation with the marine mining industry and with other stakeholders at marine mining operations. As the review of the Code is now due, to begin the process, this paper will present an analysis of the Code and initial suggestions for its modification. Keywords: marine mining, marine environment, sustainable development, marine environmental risk management

Jan 1, 2008

By S C. Dominy

The Bendigo Goldfield is located in the state of Victoria, Australia, and is one of a number of quartz-reef fields that make up the greater Central Victorian Goldfield. At Bendigo gold particles are characterised by their very coarse nature, ranging from below 100 to over 10 000 ¦m in size. Some 60 per cent of the gold occurs in particles greater than 1000 ¦m in size. This characteristic ensures that the mineralisation is very challenging to both sample and evaluate. Diamond drilling provides a measure of geological structure, but is a poor measure of in situ grade, understating it by up to 75 per cent. Drilling combined with on-reef development, bulk sampling and geological interpretation has been used to define Indicated Mineral Resources. For grade control, a balance had to be sought between sample representivity, cost, turn around time and practicality. Sampling of extreme nugget effect deposits generally requires the use of bulk sampling in order to be representative. However, the cost and turn around time is significant. As a result, a 50 kg micro-bulk sampling system has been introduced that attempts to overcome these problems by providing a fast, practical method that can discriminate between ore and waste. As the mine passes into production, grade control will increasingly become dependent upon geological data collection and interpretation.

Jan 1, 2005

Magnesium-rich olivine is used as an additive in LKABÆs iron ore pellets in order to improve the high-temperature properties of the finished products. Olivine is only added to the pellets that are manufactured for ironmaking in the blast furnace process. The olivine comes from Minelco/LKABÆs mine at Seqi in Greenland. This dunite complex of Archean age contains about 95 - 98 per cent olivine. The grain size is typically 1 - 3 mm but much larger euhedral olivine crystals are common in the porphyritic type. The serpentinisation is absent or weak. The most common minor minerals are chromite, chlorite, talc, phlogopite, amphibole and pyroxene. The crushed olivine is transported to LKABÆs pelletising plants in Northern Sweden. After grinding, small amounts of olivine are mixed with the magnetite concentrate. The concentrate is rolled into green balls, about 10 - 14 mm in diameter. The moist green balls are transferred into the pelletising machine where the maximum temperature is about 1250¦C. The residence time in the pelletising machine is about one hour. Olivine is used in order to get pellets with better high-temperature properties. This paper deals with the olivine alteration during the pelletising process and the behaviour of magnesium following the decomposition of magnetite. The reactions are quite complex in the pellets because the physico-chemical conditions vary considerably in the process from one process step to another, and in the pellet bed from the top to the bottom. Some olivine particles in the finished product show very slight alteration, some particles are completely decomposed. The formation of the iron-rich coronas around the olivine particles is the most typical alteration. In the finished pellets the magnesium is mainly distributed between the following phases: unaltered olivine, coronas, Mg/Fe-ferrite and silicate slag.

Jan 1, 2008

By R. C. Buehl, M. B. Royer

A three ton per day blast furnace using blast temperatures up to 2200°F was operated to recover manganese from open-hearth slag and manganiferous iron ore. The spiegel product containing 12 to 24 pct Mn was blown in a basic converter to high manganese oxide slag as described in the paper which follows this one. HE steel industry, which uses over 900,000 short tons of contained manganese yearly, depends on the maintenance of supplies of manganese from foreign sources such as India, Gold Coast, and Union of South Africa. Domestic ores furnish less than 10 pct of the yearly requirements of metallurgical ores and, unfortunately, the United States has no large, or even moderate, grade deposits that could supply a major portion of the requirements for many years. It may be necessary to obtain manganese from low grade materials, even though the cost is appreciably higher than when imported ores are used. An American Iron and Steel Institute questionnaire prepared in 1949 revealed the startling fact that open-hearth slags contained about as much manganese as is imported annually. The quantity of slag produced each year by integrated steel plants (not including cold-melt shops) and the manganese contained in the slag follows: Flush slag, 4,500,000 net tons a year containing 408,000 net tons Mn; tap slag, 7,230,000 net tons a year containing 364,000 net tons Mn; for a total of 772,000 net tons Mn. Some of this slag is returned to blast furnaces but about 75 pct would be available for the recovery of manganese. In several decades, when many of the present sources of high manganese ores have been depleted, open-hearth slag and manganiferous iron ore may become the major source of the manganese required for steel production. In 1948, manganese-ore imports from the U.S.S.R. were 427,000 net tons or 34 pct of the foreign ore. Imports from this source were only 81,000 tons in 1949. It became evident that adequate stockpiles of manganese could not be accumulated for many years. The Bureau of Mines initiated an extensive research program on methods for increasing domestic production of manganese and entered into a cooperative agreement with the American Iron and Steel Institute for recovering manganese from open-hearth slags whereby the Institute provided additional research funds and trained operating personnel. Description of Process and Equipment Numerous pyrometallurgical and hydrometallur-gical processes and combinations of the two can be applied to the recovery of manganese from open-hearth slag and manganiferous iron ore. The process being developed at the Pittsburgh Station of the Bureau of Mines consists of: 1—Reduction of open-hearth slag in a blast furnace to recover the iron and manganese as spiegel-eisen with 12 to 24 pct Mn. 2—Selective oxidation of the manganese from the blast-furnace metal in a basic-lined converter to produce a high manganese slag with up to 60 pct Mn content, which is a synthetic ore for the production of ferromanganese. This process has the advantage over other processes in that the equipment which would be used, such as blast furnaces and converters, has been highly

Jan 1, 1953

A Discussion at a meeting of the New York Section, Dec. 2, 1914 D. H. BROWNE.-If 35 years ago we had met to discuss the subject that is before us to-night, the criticism that we must all be mad or we would not be here at all would have been perfectly applicable. No group of men would have come together a generation ago to discuss the subject of smelting ores without extraneous fuel, or largely by the oxidation of their own impurities, without incurring the imputation of being crazy. But metallurgy, like all other branches of human industry, progresses by idealism. In every generation there are old men who see visions and young men who dream dreams. Thirty-five years ago John Hollway conceived the idea that in the Rio Tinto ore the sulphur and iron, if properly oxidized, could be made to produce heat enough to carry on the smelting of the ores. He saw the vision. Twenty-five years ago Mr. Austin, of Montana, took out a patent on a furnace in which to carry out Hollway's ideas. He dreamed the dream. Twenty years ago a man working in Montana grasped the possibilities of this idea. To him has been given that inexpressible pleasure which comes from realizing the idea and making the concept concrete. Gentlemen, it gives me great pleasure to introduce to you the man who made the ideal real, Robert Sticht, of Tasmania. ROBERT C. STICHT, Queenstown, Tasmania.-I find it extremely difficult to act up, in the smallest degree, to the enconium passed upon me by Mr. Browne. It is true it has been my good fortune to be able to put some sort of method into the "madness" of 25 years ago; but I am not the originator of the idea, nor the only one who has carried it out. Mr. Browne mentioned the originator, who undoubtedly is Mr. Hollway. He also mentioned Mr. Austin, who gave the first instigation to pyrite smelting as now practiced. As against both of these men I was fortunate enough to be able to command the necessary natural resources to carry out the idea continuously in a practical way, so as to make it permanently possible.

Jan 4, 1915

By A. E. Hall, K. R. Mutama

The wind tunnel has been used extensively for mine ventilation research and continues to be a valuable tool. The advent of the powerful personal computer has reduced the use of wind tunnels because several flow problems car be simulated and investigated quickly. This paper describes the recent design and construction of a wind tunnel which is now being used for mine ventilation aerodynamic research and teaching purposes. The wind tunnel has a square cross section measuring 900 by 900 mm and is 8485 mm long from the entrance to where it joins the fan. The tunnel was designed as an open circuit suction type and is attached to a 7.5 kW variable pitch axial flow fan capable of producing over 6 m3/s of airflow. The entire length of the wind tunnel can be used for flow investigations and all activities are visible since the sides are made of plexiglass. Results are presented from the test program to show that even without honeycombs and screens air flow symmetry is achieved.

Jan 1, 1993

By R. H. Wightman, G. H. Adams

H. C. WEED*—Referring to the use of "dummy fuse" for checking the shots in chute blasting operations, I believe that an even better practice is to blast the chutes with no delay electric blasting caps. Permanent wires can be strung through the grizzly lines and the lead wires from the shots attached to these lines. Using this method, no time is lost in spitting the individual fuses. The possibility of a fused charge becoming dislodged by previous shots and falling into the chute below is eliminated and there is no chance of a man walking back into his own blast. A small hand twist blasting machine or a radio "B" battery can be used to fire the shots if the lines are fired individually. If so desired the entire area can be blasted at once by a switch at the manway or in the drift below. providing a sufficient source of electricity is available. In case a switch is used, proper precaution for locking the switch must be provided. R. H. WIGHTMAN (author's reply) —We do use electric blasting when the number of shots is large; but we have found it more satisfactory to use fuse and caps for individual shots on the grizzly.

Jan 1, 1950

The Grants register 1979-1981, edited by Roland Turner. London, MacMillan, 1980. (Available from Geonicla Publications, P.O. Box 31487, Braamfontein 2017 at R31,37; price to libraries R28,33.) Buyer's guide. Pretoria, South African Bureau of Standards, 1980. (Available free of charge from the Director General, SABS, Private Bag X191, Pretoria 0001.) SABS specification on gold alloys. World coal resources, by Bernardo F. Grossling. London, Financial Times, 1980. 160 pp. $D.S.155. Resources policy, a quarterly journal publish d in March, June, September, and December. Annua subscription: £40. Diamond jubilee 1980 The Associated Scientific and Technical Societies was established in 1920 primarily to 'promote the interests of Scientific, Professional and Technical Societies, the advancement of knowledge in Scientific and Technical subjects, and to uphold the professional status of such Societies,' and to 'represent corporately the views of Scientific, Professional and Technical Societies and the members thereof'. In this manner the strength of unity

Jan 1, 1980

THE International First-Aid and Mine-Rescue con-test, under the auspices of the U. S. Bureau of Mines, was held at Butte, Mont., on Aug. 20, 21 and 22. As usual, the first-aid elimination contests were held the first day, the finals the second, and the mine-rescue contest the third. While the number of teams entered was not as great as at some previous contests held nearer the coal-mining districts of the East, the atten-dance was much larger than usual, due to the great interest in first aid, rescue work, and accident preven-tion that has been created by the effective support and constructive policy of the Anaconda Copper Mining Co. toward them. The winner of the first-aid contest was the Great Falls Zinc Mill team of the Anaconda company, of which Fred Erickson is captain. Second place went to the Baton Rouge team of the Standard Oil Co. of Louisi-ana, and third to the Great Falls Furnace Refinery team of the Anaconda company.

Jan 9, 1928

By Strathmore R. B. Cooke, Marcus Digre

That calcium will activate quartz for flotation with anionic collectors such as soaps has been known for a number of years,12'3 and the method has been applied to the concentration of various iron ores4 5 6 as well as to other materials. Clemmer gives the permissible pH range for the activation of quartz by calcium as between 9.5 and 12.5 with an optimum value of 11.0, and Hertzog7 has determined the calcium abstraction of iron ores and of Ottawa sand, but otherwise little quantitative data are available. The purpose of the present investigation was to determine those conditions under which dilute solutions of calcium chloride will activate quartz for anionic flotation and to determine the amount of calcium ion adsorbed by quartz. Those variables, the influence of which was primary, such as calcium, sodium, and hydrogen ion concentrations, were studied. Other factors which influenced the results are discussed in the description of the experiments. Three different experimental methods were used to determine the activation of the quartz. The first method was to treat a sample of ground, deslimed quartz with solutions of calcium chloride and sodium hydroxide, and to determine the calcium adsorbed by chemical analysis of the solution. Advantages are that a direct value for the amount of calcium adsorbed is obtained, and introduction of a collector which may affect the adsorption is avoided. The difficulty is to determine the extremely small differences in the calcium concentrations before and after the abstraction. The second method was to determine the activation of the quartz by observing the contact angle obtained on pressing an air bubble against the polished surface of a piece of quartz immersed in a solution of calcium chloride, sodium hydroxide, and sodium oleate. This is the well-known "contact angle" method. The third method was to determine those conditions under which an air bubble will pick up quartz particles immersed in a solution of the desired reagents. Disadvantages common to the sec- ond and third methods are that introduction of a collector is necessary to obtain bubble contact with the mineral; very small quantities of contaminant can give erroneous results; and no information is obtainable concerning the amount of calcium adsorbed. Advantages are that tedious analyses, which at best give approximate results, are avoided, and the conditions, especially for the third method, closely simulate those of actual flotation. Although the most consistent results were obtained with the third method, this paper describes the techniques used and the results obtained from each of the three methods. Distilled water and analytical grade reagents were used throughout except where otherwise stated. Between tests all glass articles were cleaned with a chromic-sulphuric acid mixture and sodium hydroxide. Abstraction Method The raw material for the tests was Ottawa sand sized between 14 and 28 mesh. Batches of 500 g of this sand were dry ground for 1 hr in an Abbe porcelain mill with 2700 g of quartz pebbles. The batches were split down to samples weighing approximately 30 g, and the weight of each sample was adjusted to exactly this value.

Jan 1, 1950

By Xie Gang

The leaching experiment used by bacteria on the low grade sulfide copper ore which is difficult to leach was carried out. Some leaching bacteria which have good physiological and leaching activity were isolated and domesticated, and were adapted to leaching ore. The effective domesticated method was put forward. Some technological parameters such as ore acidification, potential controlling, leaching method, ore granularity, leaching acidity, leaching liquid quantity and acid consumption were obtained by column experiment and leaching ore in situ under the ore well. The results show it could leach 30.62% of copper from the low-grade sulfide copper ore within 24 weeks by column leaching method, and the bioleaching rate of copper is 22.83% by leaching in situ under the ore well in 8 months.

Jan 1, 2006

The latest Active List of Permissible Explosives and Blasting Devices was published in July, 1929, as Bureau of Mines Report of Investigations 2947, 2 under the title or "Active List of Permissible Explosives and Blasting Devices Approved Prior to June 30, 1929.' Subsequent Reports of Investigations 2253, 2958, 2962, 2972, 2988, 2992 and 3000, were issued to record the official actions taken during July, August, September, October, 1929, and January, February, and March, 1930. No changes were made during November and December, 1929, and April, 1930. The following additional changes in the list recording the actions taken during May, 1930, are now authorized. [ ]

Jan 1, 1930

E. A. Loria (Product Metallurgical Engineer, Crucible Steel Co. of America, Pittsburgh)—In this interesting paper, our introductory work was quoted. We would like to call attention to our sequel paper on the experimental determination of oxygen in cupola-melted cast iron,20 which was not mentioned. Vacuum-fusion oxygen values (as well as hydrogen and nitrogen) were reported for nine heats of cast iron melted in the Battelle 10-in. cupola under normal operating practice and under oxidizing conditions. The oxygen analyses ranged from 12 to 68 pprn compared to the author's computed range of 10 to 80 ppm. The average amount of oxygen found in our irons was about 20 pprn and changes in the silicon content of the iron from 1.32 to 2.35 pct had no consistent effect on the oxygen content of the iron. The gas determination specimens were poured in split steel molds that produced a clean pin, 3/8 in. diam and 2 in. long. Because freezing was almost instantaneous, the pins were entirely white iron (nongraphitic). In the early stages of the investigation, the pins were transferred to a mercury-filled trap system immediately after pouring. This was done to collect gas evolved between pouring and analysis. However, it was found that during storage for 4 weeks gas evolution was negligible. Because the vacuum-fusion analysis was usually completed within 4 days of pouring, pins from later heats were not stored in the mercury-trap system. We found some evidence that cast iron picks up oxygen during long storage, because of rusting. Earlier work by the British Cast Iron Research Association has shown that cast irons may be stored for a long time without significant change in their oxygen content. The practical significance of this study (and our own) would be in the improvement of cast-iron quality. Has the author investigated this aspect and reached any conclusions on the effect of oxygen on the mechanical properties of cast iron? The second phase of our study was to determine the properties of the test bars poured simultaneously with the gas analysis specimens. We realize that there may be complicating factors attendant in this procedure.21 Results from many test specimens measuring chill depth, transverse flexure and deflection strength, spiral fluidity, and sensitivity to hardness of gray irons ranging from 12 to 68 pprn oxygen showed that the lowering of transverse strength was the only significant undesirable effect of high oxygen content. A statistical study of the chill test results21 showed that the iron containing 22 to 46 pprn oxygen had forced chill depths that were 2/32 in. below the expected value from their composition, and irons containing less than 16 ppm oxygen had forced chill depths averaging 1/32 in. greater than the expected chill depth. Higher oxygen contents, within the range of 12 to 68 pprn did not increase forced chill depth. With the wedge tests, there was a good linear relationship between carbon equivalent of the irons and their chill depth. The results indicated that oxygen contents below 50 ppm in the iron did not affect chill depth. With 50 to 70 ppm oxygen in the iron, oxygen appeared to have a slight graphitizing tendency. These results are in disagreement with the common belief in gray iron foundries that "oxidized irons" produce high chill depths. It would be appreciated if the author would comment on this subject. Gustaf Ostberg (author's reply)—In Fig. 1 the legend of line I should read 2 pct C, 1 pct Si. The author wishes to thank Mr. Loria for calling attention to his later work, which was published after the present paper was concluded. The range of oxygen contents quoted seems to agree well with the author's values. The lack of response to variations in silicon content is probably due to the fact that the oxygen content in most cases was below the saturation level. The absence of temperature dependence, even in the case of saturation, is understandable if the difficulty in formation and escape of the deoxidation products is taken into account.

Jan 1, 1960

By D. Harrington

"Since its establishment in 1910 the Bureau of Mines has consistently recommended the rock-dusting of bituminous and lignitic coal mines as a preventive of widespread explosions. As with many other proposed innovations, the response to the bureau's recommendation has been slow; but during the past 10 years many coal mines have done some rock-dusting, even though probably fewer than 50 of the country's approximately 6,000 coal mines are thoroughly rock-dusted. Notwithstanding the inadequacy of the rock-dusting which has been done in most of the mines which have adopted the new system, numerous cases are at hand. indicating the well-founded basis of the bureau's claims regarding the effectiveness of rock-dust in stopping explosions. Very brief descriptions will be given of several cases in which explosions in coal mines of the United States are fairly definitely known to have been stopped by rock-dust during the past 5 or 6 years; in addition, several cases are cited in which explosions were stopped through the agency of rock-dust and moisture.A rock-dusted mine is one in which the floor, roof, and sides of every accessible working place, including haulage ways, airways, manways, and rooms and connections have an application of rock-dust - for instance, limestone - such that the incombustible content of the aggregate dust on those surfaces is over 60 per cent. The dust in these workings should be sampled at regular intervals to determine whether the incombustible matter in the coal-dust is less than the necessary minimum of 60 per cent; another application of rock-dust should be given when the incombustible percentage is less than 60. (As an example of thorough rock-dusting, see Bureau of Mines Information Circular 6578, 1932, which describes how rock-dusting is done in Alabama at a cost of 1.8 cents per ton of coal for a thorough application; Report of Investigations 3060, February, 1931, Rock-Dusting In A Pennsylvania Coal Mine; or Information Circular 6244, 1930, giving data, including costs, on rock-dusting in coal mines in New Mexico.)An explosion may be started by the ignition of gas, even when rock-dust is present; hence it is not correct to say that rock-dust will prevent an explosion of gas; however, if rock-dust has been applied to a working place which has present an explosive mixture of me-thane and air, and if an open light is used or a blown-out shot or an electric arc occurs, although there will very likely be an ignition and possibly an explosion, the rock-dust will certainly prevent ignition of the coal-dust nearby, and unless the general atmosphere has an explosive gas mixture, the rock-dust will stop propagation of the flame throughout the mine. In other words, if an explosion is started by gas or by coal-dust and the mine has been rock-dusted as defined, the rock-dust will prevent widespread extension of the explosion, unless the general atmosphere contains methane and air in explosive proportions.With these definitions clearly in mind, instances are now given where rock-dust has played a prominent part in the stopping of several explosions. No cases have been included where there was any reasonable doubt that rock-dust had taken a definite part in the limiting of the explosions."

Apr 1, 1932

By Francis R. Redfern

Newton's method of solving a nonlinear equation is employed to develop a least-squares acoustic emission source location method in which the spatial residuals of the location equations are minimized. These residuals, related to the distances from the least-squares source location perpendicular to the surfaces described by the location equations, permit location error to be estimated to first order along the spatial axes. The solutions obtained by this method are the best spatial solutions to the location equations in the least-squares sense. Also presented is a procedure by which the geometrical control of a transducer array can be evaluated using the rate of change of source location with fractional velocity change.

Jan 1, 1982