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By Robert D. Pehlke, Robert G. Blossey

The solubility of nitrogen in liquid cobalt and a number of dilute cobalt alloys Was measured in the temperature range 1550" to 1700°C at nitrogen pressures to 1 atm. At 1600°C and 1 atm nitrogen pressure, the solubility in pure cobalt is 0.0047 wt pct [3.8 cu cm (stp) per 100 g Co]. The standard free- energy change for the solution reaction 1/2 N,(g) = gout pct in Co) is iF - 10,000 + 5.24 T cal per g atom N. The alloying elements Cv, Cu, Mo, Nb, Tu, W, and V increase the nitrogen solubility, while Al, Fe, Ni, and Si decrease the nitrogen solubility. THE solubility of nitrogen in cobalt was studied previously by Sieverts, Busch and Dodd,2 and Schenck, Frohberg, and Graf.3 Sieverts investigated the solution below the melting point of cobalt, and Busch and Dodd worked with the liquid metal; however, neither study found any solubility for nitrogen in cobalt. Schenck et al, reported a solubility of 0.005 wt pct N in cobalt at 1600°C. Seagle et al. 4 reported residual nitrogen in cobalt after electron-beam melting, although no reasons were ventured for the presence of this interstitial impurity. This investigation was undertaken to clarify these inconsistencies, and to provide additional data on atomic interactions in dilute metallic solution. EXPERIMENTAL PROCEDURE This study was well-suited to the Sieverts' method, which has not changed significantly since its introduction by Sieverts. 5 The apparatus was of the same design as that described by Humbert and Elliott, 6 with a reaction bulb similar to that reported by Weinstein and Elliott. 7 The buret was not water-jacketed, but the measured gas volumes were corrected for small changes in room temperature. The melt temperature was measured by a disappear ing-filament optical pyrometer sighted through a prism onto the melt surface. Temperature calibration was made based on a measurement of the melting point of pure cobalt taken as 1495°C. The difference between the reciprocal of the observed and the true temperatures was assumed to be a constant, neglecting the changes in emissivity with small additions of alloying elements. The total error in gas-volume measurement caused by accuracy limitations in reading the mercury level in the buret, variation in ambient temperature, and variation in melt temperature was about 0.5 cu cm. This represents approximately 10 pct of the total solubility, and is a larger relative error than in previous studies6-' because of the small nitrogen solubilities encountered in the present investigation. The source and purity of the materials used are listed in Table IA. Recrystallized alumina crucibles were used throughout the investigation. Examination of the crucibles after solidification of the melts showed no reaction or penetration. Alloying of the cobalt melt was accomplished with the dropping technique described by Humbert. 6 The investigation was started using cobalt from African Metals Corp. in shot form, but this material was not of sufficient purity and did not yield consistent solubility results. Sherritt Gordon Mines Ltd. cobalt was then used throughout the investigation. In experiments involving alloying elements which could react with oxygen dissolved in the melt, the cobalt was further refined by annealing at 1400°C in a dry-hydrogen atmosphere. This treatment, described by Woolley and Pehlke, 8 markedly reduced the oxygen content. The oxygen and nitrogen analyses of the cobalt used in this study are presented in Table IB. RESULTS Hot Volume. The hot volume was measured prior to each run and was 50 to 55 cu cm (stp) at 1600°C. It increased linearly with increasing pressure and decreasing melt temperature. Pure Cobalt. Measurements of the solubility of nitrogen in pure cobalt and the temperature coefficient of solubility are summarized in Table 11. At 1600°C and 1 atm nitrogen pressure, the solubility was found

Jan 1, 1967

By I. W. Alcorn

THE cost of drilling in the past few years of proration and ever decreasing allowables has received increasing thought and study. It seems to parallel the strides made with respect to pumping problems in 1932 and 1933. Low profit margins stimulate the search for economies. Any discussion of slim-hole drilling involves a multitude of abstract values, and the problem can be attacked only by a study of records and rational thinking. In a measure, traditional methods must be studied with a questioning attitude. The words "slim hole" may be somewhat misleading, as the term is entirely relative. For that reason I wish to define it as used in this discussion. On the Gulf Coast, I would define a normal hole as one in which 10 3/4-in. surface casing has been set and below which a 9 7/8-in. hole is dug. With this as a datum, holes of larger diameter below surface would be called "large holes," with 12 1/4 in. as the "extra large" class and holes of smaller diameter classed in the slim-hole range, 6 1/4 in. representing about the minimum size. The optimum size has not been definitely determined, but the preference is for a program that would call for setting 8 5/8-in. surface casing and digging a 7 3/4-in. hole. Different sections of the country might very properly deviate from this practice because of differences in depth, formations, regulations, or other factors. This paper will be confined to a discussion of the drilling of exploratory wells, as the drilling of this class of wells is quite a different problem from that of drilling a proven field, although many of the same considerations apply in both cases, as will be apparent. Although for many years shallow slim-hole and core-test holes of small diameter had been dug, it was not, I believe, until the fall of 1937 that serious consideration was given to its application in drilling wells 7000 ft., 8000 ft. and 9000 ft. deep on the Gulf Coast. This does not mean that nearly all such wells drilled on the Gulf Coast now are of the slim-hole variety, but only that efforts have been aimed in that direction with very encouraging results and there has been a very definite trend to reduce the size of holes, a trend that probably will continue. As to exploration policy, there appear to be two distinct and opposite views. The one usually followed in the past was more or less designed to "shoot the works" on one well. Generally on the Gulf Coast, the depth at which production might be obtained is limited by the drilling equipment and economics. With this thought in mind, wildcat wells are often planned with the idea of carrying the depth to 10,000 ft., or even 15,000 ft. All of the possible hazards are anticipated and the general drilling and casing programs are so predicated. Such wells contemplate the drilling of large-diameter holes; the setting of at least one protective string of casing and an extensive coring policy, all of which finally adds up to an expensive well. As opposed to this general policy is the slim hole. This, on the Gulf Coast, recognizes that new discoveries represent gam-

Jan 1, 1941

By AIME AIME

As already announced in the January Bulletin, this publication will be issued during the coming year monthly instead of bi-monthly as heretofore. Among other reasons for this change, it is desired to effect thereby the earlier transmission to members of important papers, lists of candidates for membership, and other items of timely information. Spokane Meeting and Excursions. Further details of the 97th meeting of the Institute have been sent to members in the Circular of May 8, 1909. The list of those intending to go on the excursion is filling up, and since preference in position of state-room or berth is given in the order of application, members and guests who expect to accompany the party and have not yet communicated with Mr.. Theodore Dwight, 29 West 39th St., New York, N. Y., are earnestly requested to do so without delay. For convenience of the members a summary of the additional information is here presented. A special train of Pullman cars will leave Chicago about September 15. This train will be side-tracked at all places visited, and the sleepers will be available for lodging and the dining-car for meals during the entire trip, except during the Yellowstone Park tour. The general cost of the trip per member will be $300, which includes transportation, berth, and meals for the entire trip from Chicago back to Chicago, almost 6,000 miles, and the tour through. Yellowstone Park, occupying in all about 30 days. Special, accommodations on the train will be furnished at the following rates

Jun 1, 1909

By P. Rama Rao, D. H. Sastry, T. R. Anantharaman

Deformation (a) and twin (a') stacking-fault densities in cold - wor ked filings of fcc silver alloys containing 2, 4, and 6 wt pct Sb have been estimated from Debye-Scherrer peak shifts and asymmetries, respectively. Evidence has been obtained for a possible strain-induced transformation in a two-phase alloy containing 9.0 zut pct Sb. Using Fourier analysis of X-ray line profiles, the effective particle sizes and root mean square strains have been determined for the three fcc alloys in the cold-worked condition. The values of the compound fault probability 0.5a + a') evaluated from peak shifts and asymmetries are in reasonable apeement with those derived from peak broadening. COLD work of fcc metals may produce both deformation and twin (also called growth) stacking faults, which can be detected and quantitatively estimated by X-ray diffraction techniques.' In addition to peak broadening, deformation faults produce a shift in peak positions and twin faults an asymmetry of the Debye-Scherrer reflections. Besides stacking faults, a reduction in the size of coherently diffracting domains and distortion within these also broaden the powder reflections of deformed fcc metals and alloys. Fourier analysis' of peak shapes provides quantitative infor- mation about the effective domain sizes, lattice strains, and also the compound fault probability (1.5a + a'), where a and a' represent deformation-and growth-fault density, respectively. The Paterson method2 for determination of a from peak shifts, the warren3 or wagner4 method for evaluation of a' from peak asymmetries, and the Fourier analysis for quantitative estimates of all the contributory causes of peak broadening have, in the course of the last decade, been extensively applied to obtain data for a number of deformed fcc metals and alloys. One of the useful results of earlier measurements of a in several solid solutions is a correlation between a and the solute valence as well as electron concentration per atom (p) in alloy systems based on a particular solvent. The work of Smallman and westmacott5 on Cu-Zn, Cu-A1, and Cu-Sn, of Foley, Cahn, and ~aynor' on Cu-Si and Cu-Ge, and of Rama Rao and Anantharaman7 on Cu-Si and Cu-Sb has revealed a common trend in the increase of a with p and also a with valency of the solute. Cahn and Davies' and Ad-ler and wagnerg have observed similar trends for silver alloys containing cadmium, indium, and tin. Data on a for silver alloys with a pentavalent solute element are not, however, available. It is primarily to fill this need that the present investigation on Ag-Sb alloys was undertaken. Utilizing the Patersonand Warren methods, values of a and a', respectively, have been determined in three silver alloys containing 2, 4, and 6 wt pct Sb. By Fourier analysis of the broadening of Debye-Scherrer peaks, effective particle sizes, lattice strains, and the compound fault probability (1.5a +a') have also been evaluated.

Jan 1, 1967

By Luther B. Eames

J. V. N. DORR, New York, N. Y.-I have read Mr. Eames' excellent paper on this' subject with great interest, for besides being connected with the design and installation of the first modern countercurrent decantation plant, he was associated with me in the development of the thickener in the Black Hills some years earlier, and has now been operating the largest plant that has used this method of treatment. His, discussion, of the influence of different factors on the results obtained is quite thorough, but several features I think may not be apparent to the reader who has not studied the subject carefully. 1. Grade of Ore That Can Be Treated.-Mr. Eames has shown graphically that' the assay of all solutions increases directly as the grade of the ore treated. He has not emphasized, however, the fact that, as most operators prefer to keep the assay of pregnant solutions down below a certain figure in order to insure complete precipitation, an increase in ore value beyond a certain point means a proportional increase in solution precipitated and therefore not only an increase in the percentage of dissolved gold recovered, but an actual descrease in tailing loss. The importance of the influence of the percentage of moisture in the underflow on loss is made clear, and I. may add that unless it is quite evident that pulp can be safely settled to 50 per cent. solids, or more, it is usually advisable to use a continuous filter at the end of. a decantation plant to reduce losses in cyanide and gold. Dissolution in Thickeners.-The influence of a change of solution on silver dissolution has long been emphasized and most gold operators have. recognized the annoying way in which a gold ore after being agitated until nothing more can be dissolved will loosen up again when almost ready for the dump after the dissolved metal has been removed by decantation or filtration. Continuous decantation gives favorable conditions for this additional dissolution and a chance to recover most of what is thus dissolved. This is especially true if an extra tank is used in the series and barren solution added at the second tank from the end instead, of the first, as shown in Mr. Eames' Fig. 1. It can be calculated that the use of this tank will reduce the mechanical loss of cyanide by one-third and cut the gold loss as well; but the saving in gold or silver that may dissolve when the pulp is first diluted with weak solution may be as important.

Jan 4, 1917

By T. F. Kassner

The dissolution kinetics of type 304 stainless steel in the Bi-Sn eutectic alloy have been investigated under the well-defined hydrodynamic conditions produced by the rotating-disc sample geometry. In addition, the mutual solubilities of iron, chromium, nickel, and manganese from 304 stainless steel in the eutectic alloy were determined over the temperature range 450" to 985°C. The convective -diffusion model for mass transport from a rotating disc was used to interpret the experinlental dissolution data. The dissolution process was found to be liquid-diffusion-controlled under specific conditions of temperature and Reynolds number. Liquid penetration into the 304 stainless steel resulted in a reduction of the di,ffusion-controlled mass flux and thus precluded the calculation of the diffusion coeficients of the four components from 304 stainless steel in the Bi-Sn eutectic alloy. The convective-diffusion model for diffusional limitations of electrode reactions and mass transport at the tationssurface of a rotating disc set forth by Levich 1,2 has found wide applicability in the investigation of electrochemical and dissolution phenomena in aqueous systems. Riddiford 3 and Rosner have reviewed the model and also include numerous references on work of this nature. More recently the rotating-disc system has been applied to the investigation of hetereogeneous reactions in liquid-metal systems. Shurygin and Kryuk 5 have measured the dissolution rates of carbon discs in molten Fe-C, Fe-Si, Fe-P, and Fe-Ni alloys. Shurygin and shantarin6 also studied the dissolution kinetics of iron, molybdenum, chromium, and tungsten, and the carbides of chromium and tungsten in Fe-C solutions with a rotating-disc sample geometry. In these systems it was possible to distinguish between diffusion and reaction control mainly through experimental confirmation of the velocity dependence of the dissolution rate predicted by the model. However in the absence of dependable solubility data and the virtual lack of diffusion data in these systems, a quantitative check of the magnitude and the temperature dependence of the rate was not possible. In many instances, estimates of the activation energy for solute diffusion and the diffusion coefficient based upon the experimental dissolution data are not credible. A recent study by this author7 has resulted in a critical test of the model in a liquid-metal system. The solution rates of tantalum discs in liquid tin were measured over a wide range of temperature and velocity conditions. In addition, the solubility and diffusion coefficient of tantalum in liquid tin were determined as a function of temperature. The latter data were used with the model to predict both the magnitude and the temperature dependence of the dissolution flux. In that work it was also deemed necessary to reevaluate the solution to the convective diffusion equation to incorporate the effect of the lower range of Schmidt numbers encountered in liquid-metal systems. Good agreement between the model and the experimental dissolution data in the region of diffusion control was obtained in the Ta-Sn system. The Bi-Sn eutectic alloy is used as a seal between the reactor head and the reactor vessel in the Experimental Breeder Reactor-11. The alloy is fused periodically prior to fuel-handling operations. In that connection, it was necessary to investigate the compatibility of the liquid alloy with the type 304 stainless-steel containment material. The results of a rotating-disc study in this multicomponent system are presented. EXPERIMENTAL METHOD The 5.08-cm-diam discs were machined from 0.317-cm-thick plate. Chemical analysis information for the type 304 SS material is given in Table I. The discs were ground flat on metallographic paper and given a final polish on Linde B abrasive. A thin support rod was threaded into the disc and the region around the threads was fused under an inert gas. The support rod was fitted with a quartz protection tube and then was attached to a supporting shaft which passed through a rotary push-pull vacuum seal. The disc and supporting shafts were dynamically balanced prior to insertion into the furnace tube. The apparatus is shown schematically in Fig. 1. The 58 pct Bi-42 pct Sn eutectic alloy melts were prepared from 99.995 pct pure Bi and Sn by fusing the components in a 7-cm-ID Pyrex crucible. The system in which the melts were made was evacuated to a pressure of 1 x 10-6 Torr and back-filled with purified argon several times before melting the charge. The ingot was reweighed and placed in a slightly larger-diameter Vycor crucible used in the dissolution runs. A run was started by lowering the disc into the liquid

Jan 1, 1968

Coal Division Industrial Minerals Division Minerals Beneficiation Division Mining, Geology and Geophysics Division OFFICERS AND COMMITTEES S D Michaelson, President J W Woomer, President-elect Elmer A Jones, Past President C E Lawall, Eastern Regional Vice-President G Donald Emigh, Central Regional Vice-President J D Forrester, Western Regional Vice-president R B Ladoo, Treasurer John Cameron Fox, Secretary

Jan 1, 1958

By A. H. Underwood, M. –Ls. Trépanier

INTRODUCTION The King and Beaver mines of Asbestos Corporation Limited (ACL) were joined in 1957 to form the King-Beaver mine. In 1964, ACL purchased the Johnson's Asbestos Company. The combined underground workings are a part of the King- Beaver mine. Asbestos ore is mined at the rate of 4080 metric tons per day (mtpd) to provide 15% of the company's daily requirement in the Thetford Mines area. Figure 1 shows the location of Thetford Mines. Two shafts service the underground. No. 3 shaft, 457 m deep, is used for personnel and material transportation, downcast ventilation, and ore hoisting. No. 4 shaft, 820 m to the south, is 389 m deep. It is used for material transportation and upcast ventilation. Figure 2 shows the relative locations of the shafts and the underground workings. The ore is mined by block caving. HISTORY The orebody outcropped at surface and was first mined as an open pit. Underground mining commenced when open pit mining became impractical. Shrinkage stoping, used initially, was unsatisfactory in the highly fractured ore. Sublevel stoping improved ore recovery and safety, but did not overcome the adverse effects of the un- stable rock. Observations of the rock failures indicated that block caving would be satisfactory. This was confirmed by careful investigation and testing. A method of block caving similar to that used at the Miami Copper Company was adopted. The first block commenced production in June 1933. BLOCK CAVING General A large block of ore is undercut above pre- pared drawpoints. The ore above the undercut caves by its own weight. Caving progresses up- ward as the caved ore is drawn. A large proportion of the rock breaks up in the caving action so that it will pass through the drawpoints. Oversize blocks are broken by drilling and blasting or mudcapping. Figure 3 shows the progressive stages of block caving. In the central part of the ore zone, the vertical dimension of a block is determined by the 132 metres between production levels. When isolated ore pockets are developed, the vertical dimensions of a block depend on the extension of the ore zone above the undercut. The minimum horizontal dimension of an under- cut determines the cavability of the overlying block. When necessary, caving is assisted by drilling long horizontal holes into the block above the undercut and blasting them. This procedure speeds up or restarts caving. It will not ensure complete caving of blocks having undercuts less than 30 m wide. Orebodies suitable for block caving are highly fractured and wide enough to cave freely without excessive dilution. Undercutting must

Jan 1, 1981

ALASKA Established November 19, 1947 Meets fourth Monday of each month at Fair¬ banks, Alaska J D CRAWFORD, Chairman BRUCE I, THOMAS, Vice-Chairman C D JERMAIN, Vice-Chairman for Southeastern Alaska EARL H BEISTLINE, Secretary-Treasurer Box 265, College, Alaska

Jan 1, 1952

By I. W. Alcorn

The cost of drilling in the past few years of proration and ever decreasing allowables has received increasing thought and study. It seems to parallel the strides made with respect to pumping problems in 1932 and 1933. Low profit margins stimulate the search for economies. Any discussion of slim-hole drilling involves a multitude of abstract values, and the problem can be attacked only by a study of records and rational thinking. In a measure, traditional methods must be studied with a questioning attitude. The words "slim hole" may be somewhat misleading, as the term is entirely relative. For that reason I wish to define it as used in this discussion. On the Gulf Coast, I would define a normal hole as one in which 10 3/4-in. surface casing has been set and below which a 9 7/8-in. hole is dug. With this as a datum, holes of larger diameter below surface would be called "large holes," with 121/4 in. as the "extra large" class and holes of smaller diameter classed in the slim-hole range, 6 1/4 in. representing about the minimum size. The optimum size has not been definitely determined, but the preference is for a program that would call for setting 8 5/8 -in. surface casing and digging a 7 3/4 - in. hole. Different sections of the country might very properly deviate from this practice because of differences in depth, formations, regulations, or other factors. This paper will be confined to a discussion of the drilling of exploratory wells, as the drilling of this class of wells is quite a different problem from that of drilling a proven field, although many of the same considerations apply in both cases, as will be apparent. Although for many years shallow slim-hole and core-test holes of small diameter had been dug, it was not, I believe, until the fall of 1937 that serious consideration was given to its application in drilling wells 7000 ft., 8000 ft. and 9000 ft. deep on the Gulf Coast. This does not mean that nearly all such wells drilled on the Gulf Coast now are of the slim-hole variety, but only that efforts have been aimed in that direction with very encouraging results and there has been a very definite trend to reduce the size of holes, a trend that probably will continue. As to exploration policy, there appear to be two distinct and opposite views. The one usually followed in the past was more or less designed to "shoot the works" on one well. Generally on the Gulf Coast, the depth at which production might be obtained is limited by the drilling equipment and economics. With this thought in mind, wildcat wells are often planned with the idea of carrying the depth to 10,000 It., or even 15,000 ft. All of the possible hazards are anticipated and the general drilling and casing programs are so predicated. Such wells contemplate the drilling of large-diameter holes; the setting of at least one protective string of casing and an extensive coring policy, all of which finally adds up to an expensive well. AS opposed to this general policy is the slim hole. This, on the Gulf Coast, recognizes that new discoveries represent gam-

Jan 1, 1941

By I. W. Alcorn

The cost of drilling in the past few years of proration and ever decreasing allowables has received increasing thought and study. It seems to parallel the strides made with respect to pumping problems in 1932 and 1933. Low profit margins stimulate the search for economies. Any discussion of slim-hole drilling involves a multitude of abstract values, and the problem can be attacked only by a study of records and rational thinking. In a measure, traditional methods must be studied with a questioning attitude. The words "slim hole" may be somewhat misleading, as the term is entirely relative. For that reason I wish to define it as used in this discussion. On the Gulf Coast, I would define a normal hole as one in which 10 3/4-in. surface casing has been set and below which a 9 7/8-in. hole is dug. With this as a datum, holes of larger diameter below surface would be called "large holes," with 121/4 in. as the "extra large" class and holes of smaller diameter classed in the slim-hole range, 6 1/4 in. representing about the minimum size. The optimum size has not been definitely determined, but the preference is for a program that would call for setting 8 5/8 -in. surface casing and digging a 7 3/4 - in. hole. Different sections of the country might very properly deviate from this practice because of differences in depth, formations, regulations, or other factors. This paper will be confined to a discussion of the drilling of exploratory wells, as the drilling of this class of wells is quite a different problem from that of drilling a proven field, although many of the same considerations apply in both cases, as will be apparent. Although for many years shallow slim-hole and core-test holes of small diameter had been dug, it was not, I believe, until the fall of 1937 that serious consideration was given to its application in drilling wells 7000 ft., 8000 ft. and 9000 ft. deep on the Gulf Coast. This does not mean that nearly all such wells drilled on the Gulf Coast now are of the slim-hole variety, but only that efforts have been aimed in that direction with very encouraging results and there has been a very definite trend to reduce the size of holes, a trend that probably will continue. As to exploration policy, there appear to be two distinct and opposite views. The one usually followed in the past was more or less designed to "shoot the works" on one well. Generally on the Gulf Coast, the depth at which production might be obtained is limited by the drilling equipment and economics. With this thought in mind, wildcat wells are often planned with the idea of carrying the depth to 10,000 It., or even 15,000 ft. All of the possible hazards are anticipated and the general drilling and casing programs are so predicated. Such wells contemplate the drilling of large-diameter holes; the setting of at least one protective string of casing and an extensive coring policy, all of which finally adds up to an expensive well. AS opposed to this general policy is the slim hole. This, on the Gulf Coast, recognizes that new discoveries represent gam-

Jan 1, 1941

By Persifor Frazer

The subject of Egypt, to use the words of perhaps the second of modern writers on the subject [Deodat. de Dolomieu, in Observations sur la Physique, etc., January, 1793, vol. xlii., pp. 41+, 108+; Abbe Rozier, J. A. Mongez, and J. C. Delametherie, Paris], is one which strongly excites our interest, and everything relating to it demands our respect. "Thus," he continues, "the Greeks appeared to the Romans," ..." thus Egypt appeared to the Greeks, and thus all three appear to us." This writer, in the three parts of the memoire given above, speaks of Norden as the modern pioneer of Egypt, and does not seem to have made original observations himself; nevertheless he seems to have rendered the same service to his authority that Playfair not very long afterward rendered Hutton. The limitations of a paper will permit little else than citations, but these are so numerous that, unless much compressed, they would supply a chapter of moderate size. A great deal of the very interesting memoire of Dolomieu is taken up with Quixotic tilts

Jan 1, 1883

By R. D. Satterley

The Sherwood mine is an iron-ore mine owned and operated by the Inland Steel CO. in the Iron River district of the Menominee Range in Michigan. The property consists of an 80-acre tract in the village of Mineral Hills, about a mile north of the city of Iron River. The Ore mined is a reddish brown, nonbessemer hematite and limonite. The ore bodies lie in a large syncline having an east-west strike and are associated with a rock structure that is folded in a complex manner. The two 40-acre tracts run north and south and the shaft is at about the center of the north forty. The shaft is 1286 ft. deep, with the principal operating level at a depth of 1000 ft. from surface. A new level at a depth of 1200 ft. is in process of development. The ore body at the 1000 ft. level is east and west at approximately the line between the north and south forties and at this elevation is from 50 to 250 ft. wide and extends above the level at varying heights of from 50 to 250 feet. The 1000-ft. haulage drift extends south from the shaft, through the east-west ore body to the slates on the south side of the body. The ore body was intercepted in the drifting at about goo ft. from the shaft and was about 200 ft. wide at that point. A crosscut was extended to the east and west property lines along the south boundary of the ore body. This east-west ore body is laid out in areas go ft. wide for stoping with pillars 40 ft. wide. With this plan a stope is approximately 80 by 200 ft. and has a pillar on both sides that is approximately 40 by 200 ft. The ore has a height of from 50 to 250 ft. above the elevation of the level, therefore the height of the stopes varies in accordance with the height of the ore body at each location. From the crosscut, two main raises, each 6 by 7 ft., are put up into each 80 ft. stope area (Fig. 1). One of the main raises serves as a manway and the other as an ore raise, through which the ore is transferred from the stope to the tramming equipment. The original layout as for scraping of the ore from under the stopes into the ore raises. The ore raise is put halfway (east and west) along the 80-ft. stope area and usually is in iron formation. The manway and ore raise are on approximately 15-ft. centers. These main raises are Put up to a height of from 22 to 30 ft. above the level floor, then from the toll of the ore raise a scraper drift (8 X 8 ft.) is driven across the width of the ore body, which is in the neighborhood of 200 ft. From the scraper drift, mill raises are put up at 12½-ft. centers, staggered on opposite sides of the drift, and these are extended 23 ft. above the elevation of the scraper-drift floor. The sublevel connecting the top of these mill raises is called a mill "sub." Other sublevels are developed parallel to the scraper and mill subs at 25-ft. vertical intervals above the mill sub. The mill sub and the other subs above the mill sub are known as sublevels or dog drifts, and are 4 ft. wide and 6 ft. high. These dog drifts

Jan 1, 1946

By R. D. Satterley

THE Sherwood mine is an iron-ore mine owned and operated by the Inland Steel Co. in the Iron River district of the Menominee Range in Michigan. The property consists of an 80-acre tract in the village of Mineral Hills, about a mile north of the city of Iron River. The ore mined is a reddish brown, nonbessemer hematite and limonite. The ore bodies lie in a large syncline having an east-west strike and are associated with a rock structure that is folded in a complex manner. The two 40-acre tracts run north and south and the shaft is at about the center of the north forty. The shaft is 1286 ft. deep, with the principal operating level at a depth of 1000 ft. from surface. A new level at a depth of 1200 ft. is in process of development. The ore body at the 1000 ft. level is east and west at approximately the line between the north and south forties and at this elevation is from 50 to 250 ft. wide and extends above the level at varying heights of from 50 to 250 feet. The 1000-ft. haulage drift extends south from the shaft, through the east-west ore body to the slates on the south side of the body. The ore body was intercepted in the drifting at about 800 ft. from the shaft and was about 200 ft. wide at that point. A crosscut was extended to the east and west property lines along the south boundary of the ore body. This east-west ore body is laid out in areas 80 ft. wide for stoping with pillars 40 ft. wide. With this plan a stope is approximately 80 by 200 ft. and has a pillar on both sides that is approximately 40 by 200 ft. The ore has a height of from 50 to 250 ft. above the elevation of the level, therefore the height of the stopes varies in accordance with the height of the ore body at each location. From the crosscut, two main raises, each 6 by 7 ft., are put up into each 80 ft. stope area (Fig. I). One of the main raises serves as a manway and the other as an ore raise, through which the ore is transferred from the stope to the tramming equipment. The original layout was for scraping of the ore from under the stopes into the ore raises. The ore raise is put halfway (east and west) along the 80-ft. stope area and usually is in iron formation. The manway and ore raise are on approximately 15-ft- centers. These main raises are put up to a height of from 22 to 30 ft. above the level floor, then from the top of the ore raise a scraper drift (8 X 8 ft.) is driven across the width of the ore body, which is in the neighborhood of 200 ft. From the scraper drift, mill raises are put up at 12 ½ -ft. centers, staggered on opposite sides of the drift, and these are extended 23 ft. above the elevation of the scraper-drift floor. The sublevel connecting the top of these mill raises is called a mill "sub." Other sublevels are developed parallel to the scraper and mill subs at 25-ft. vertical intervals above the mill sub. The mill sub and the other subs above the mill sub are known as sublevels or dog drifts, and are 4 ft. wide and 6 ft. high. These dog drifts

Jan 1, 1945

By J. D. Cooner

SINCE my working life of 32 yr has been spent in and about the anthracite mines of the Hudson Coal Co., and the previous 4 yr in a college school of mines, I can write best about the safety program of a coal mining company. However, most of the suggestions, statements and practices mentioned herein are applicable to the members of a safety organization and program of almost any industry. It is realized that industrial officials actually interested in real safety work are fully acquainted with the requirements of a practical safety program. Therefore, for such persons there may be nothing new in this paper. The Hudson Coal Co. of Scranton, Pa., employs 7,000 men and produces 5,000,000 tons of anthracite annually at nine collieries scattered for a distance of 40 miles throughout the Lackawanna and Wyoming valleys of northeastern Pennsylvania. For the past 17 yr of the above mentioned 32, I have been engaged in safety work as a safety inspector, assistant mine superintendent, and at present the safety engineer. I have also had the privilege of visiting coal and metal mines in other parts of the United States and Canada. We Americans are the most careless people in the world. If any one doubts this statement why are 100,000 of our people killed accidentally annually? Or why during the almost four years we were engaged in World War II were more of our people killed accidentally at home than were killed in our army, navy, marine corps, coast guard, and other services during the same period? We are careless everywhere in our homes, on the highway, at work and at play. The purpose of this paper is to outline the requisites of a safety department and its program to stop the accidents occurring to employees at work in mines or industrial plants. The first requisite for a safety program is that the management must be convinced of the need and value of a real safety program. By "management" is meant each department head from the president down. Otherwise a good safety department with a good safety program will not succeed because it will not have the proper personnel and equipment to work with, its rules and regulations will not be enforced, and the members of the safety department will be belittled in the eyes of most of the workmen and supervisors. Next, a safety engineer and his inspectors should have worked in the production end of the mines as workmen and also in an official capacity. They are then in a position to have practical and first hand knowledge of the accident hazards of the industry, how to prevent them, and in general the problems of both the workmen and officials. Some men are appointed as safety inspectors when they are physically unable to carry on as mine officials. Such men may know the safety end of the industry and even be safety-conscious, but if they are physically unable to be mine officials they are also physically unable to be safety inspectors, as a safety inspector in our mines must walk from 2 to 4 miles on each daily inspection-and visit each working

Jan 1, 1948

By Evert Hoek

Summarising a conference is never an easy task, but I will try to keep my comments brief and to the point. There are a number of lessons which we have learned from our deliberations together during these past three days, and I will try to give you my impressions of these lessons and of their practical implications in open pit mining. The first thing that has become increasingly obvious during these discussions is that it is no longer acceptable to offer an 'Act of God' as an excuse for a major slope or tailings dam failure. Disasters such as that at Aberfan in the United Kingdom have focussed the world's attention on the activities of the mining industry and the report of the commission set up to investigate the Aberfan disaster clearly showed that adequate knowledge was available and that failure to apply this knowledge did not qualify as an 'Act of God'. The courts and the world's press are certain to impose increasingly harsher penalties and criticism on those held responsible for major failures, particularly of tailings dams, in the future. One of the major problems in applying available knowledge to slope problems is that attempts to apply this knowledge are frequently left too late. To provide an engineering solution in the case of a slope in an advanced state of failure is an almost impossible task, and it is essential that those responsible for the stability of a slope must learn to recognise signs of distress at an early enough stage for effective remedial measures to be taken. As our ability to solve slope problems improves, the time at which slope stability problems are considered in a mining project must move forward to a stage at which these problems can be anticipated and dealt with before they arise. Conferences such as this are important in that they enable us to exchange ideas and advance our knowledge to a level at which such predictive engineering will be possible. It has been encouraging to hear of mining projects such as the Bougainville Copper operation where slope stability studies were carried out as part of the overall evaluation programme and it is hoped that this will become a more common procedure in the future. Providing a safe slope in an opencast mine obviously costs money, either in terms of the reduced ore to waste ratio given by flat slopes or because of the cost of stabilization techniques such as drainage or the use of rock anchors. An optimum slope design will provide maximum safety at minimum cost but achieving this optimum is not an easy task. Any one of us could design a slope which would be stable to the end of time, but it is probable that that slope would be so flat as to be totally uneconomic in most open pit mining situations. On the other hand, if the overall economics of a mining project are so critical as to require slope angles which the engineer responsible for these slopes considers to be unsafe, it is his duty to make us could design a slope which would be stable to the end of time, but it is probable that that slope would be so flat as to be totally uneconomic in most open pit mining situations. On the other hand, if the overall economics of a mining project are so critical as to require slope angles which the engineer responsible for these slopes considers to be unsafe, it is his duty to make

Jan 1, 1971

Jan 1, 1933

[Billings Petroleum Section: Fraser M. Burback, chairman; N. J. Matthews, first vice-chairman; Charles B. Evans, second vice-chairman; and lion M. Madden, secretary-treasurer. Meets 3d Tues., Petroleum Club. California Coastal Section: Jack R. Severns, chairman.; Gilbert L. Jemmott, vice-chairman,- Jack A. Biren, secretary; and Robert L. Richardson, treasurer. Meets 3d Thurs.] [ ]

Jan 1, 1961

By Robert Archer

THE proper material of construction for a given purpose is that material which meets the requirements satisfactorily at the lowest ultimate cost. It is consistent with this principle that most aluminum castings are made of an alloy containing about 8 per cent. copper. This alloy possesses mechanical properties satisfactory for general engineering purposes together with manufacturing characteristics leading to the production of finished castings at low cost. It is to the aluminum industry what gray iron is to the iron and steel industry. Just as some services require "semi-steel," malleable castings or steel castings instead of gray iron, there is a certain and growing demand for aluminum castings of superior strength or ductility, or both. Since the beginning of the industry producers and others have sought to develop alloys or processes by which such castings could be made. Out of these efforts have emerged many alloy compositions and special processes, of which only a few have had sufficient merit to survive. It is the purpose of this paper to describe those which seem to the authors to offer the greatest promise for the production on a commercial scale of aluminum castings of superior mechanical properties, with special reference to heat-treated castings. The suitability of an alloy or process for actual manufacture is affected by so many factors that the problem of appraisal is not a simple one. Since this is such an important part of the whole subject it is proposed to discuss briefly some of the requirements. The final verdict is of course the result of commercial trial and experience. It will therefore, perhaps, be profitable to employ a somewhat historical method of treatment in discussing the subject of alloy development.

Jan 9, 1926

By A. M. Gaudin

The study that is presented here was undertaken in order to condense information concerning comminution, and covering a great variety of conditions, into one or several rules which would be of use in the development of a systematic theory as well as in many directly practical ways. As the result of this work, the following conclusions appear justified: 1. Rocks may be classified, from the point of view of comminution, into two general classes which are (a) structurally homogeneous, and (b) structurally heterogenous. Structurally homogeneous rocks are those in which fracture takes place through grains and grain boundaries alike; heterogeneous rocks are those in which fracture passes in preference through grain boundaries. 2. If a sized product composed of structurally homogeneous rock grains is crushed, a product results in which the relation between the percentage weight of grains of various sizes and these sizes themselves follows a definite law. This is not true of structurally heterogeneous rocks. 3. Ball milling with structurally homogeneous rocks, and under certain conditions, presents exceptions to this rule. 4. In ball milling with structurally homogeneous rocks, there exists a critical ratio of size of feed to size of balls, above which the machine operates essentially by surface wear, and below which it operates essentially by breaking across the body of the particle. 5. In ball milling, with structurally homogeneous rocks, and when the size of feed grains exceeds the critical size, the particles worn off are of a substantially definite size which depends upon the ball size. 6. In the case of a typical heterogeneous rock, such as Manhattan schist, particles are crushed in such a way that there is a greater proportion of grains of the size that is the average grain-size of the rock than there would be in that range of size with a structurally homogeneous rock. This has been found to be true with all the crushers tested. 7. Different methods of crushing produce grains of different shapes. 8. The importance of sizing the — 200-mesh portion of a crushed product, in crushing-efficiency investigations, has been greatly under-

Jan 1, 1926