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By P. D. Southgate

VIBRATION, both ultrasonic and sonic, can affect the course of a variety of metallurgical processes. Reports of work on this subject have appeared at intervals over the last twenty years and a thorough review has been made by Hiedemann,' who gives a full list of references. It has been established that both refinement of structure and outgassing of the melt can be produced in most metals with sufficiently intense vibration. Many of the experiments in this field have been carried out using alloys of direct commercial interest in production- type casting equipment. Here a different approach is adopted, and simple binary alloys which exhibit different types of structure are used to facilitate examination of the effects produced under well-controlled conditions. Results from this kind of experiment should enable the basis of larger scale experiments with similar types of materials to be better substantiated. The alloys selected for this work are: 1) aluminum of 99.2 pct purity, the principal impurity being iron, near 0.4 pct; 2) A1-1.8 pct Fe, which is near eutectic composition; 3) A1-4 pct Fe—the second phase, of composition FeAl3, forms distinct acicular crystals; and 4) A1-12 pct Si—although this is of eutectic composition, the silicon normally appears in the form of large plates. Experimental Method There are three practicable methods by which acoustic vibration can be transferred to liquid metals: a) vibration of the whole of the container; b) electromagnetic induction; and c) insertion of a vibrating probe into the liquid. Method c is chosen here, and a frequency of 8 kc per sec is used. The end of the probe is electrically heated so that the metal near it will be the last to solidify. This ensures that vibration is transmitted directly to the liquid metal without passing through any solidifying and pasty intermediate region, which would attenuate acoustic waves severely. In addition, the heating prevents solid metal adhering to the probe, with consequent detuning. The transducer assembly is shown in Fig. 1. A stack of Permendur laminations 30 cm long acts as a half-wave vibrator and a winding along its length carries polarizing and energizing currents. It is joined by a half-wave coupling bar to a mild steel exponential horn, which tapers from 3.8 to 1.1 cm diam. A stainless steel stub 3 cm in length is welded on the free end. The assembly is supported by a plate located at the node of the coupling bar. Acoustic radiation losses are minimized by allowing the free end of the transducer to protrude above the surface of the cooling water. The exponential horn serves the function of matching the acoustic output impedance of the transducer to the radiation resistance of the probe into the melt. Electrical drive is provided by a 1 kw generator. Measurements of probe vibration amplitude during melt treatment are obtained from a gramophone pick-up with the needle attached to the wide end of the horn. This device is calibrated with a microscope before use. Liquid aluminum is very actively erosive and solution of materials with which the molten metal comes into contactcauses difficulties. Erosion of the probes in particular can be severe since it is greatly accelerated by the vibration. Tests of various acous-

Jan 1, 1958

By Edward C. J. Urban

THE nature of certain pernicious dusts commonly encountered in the removal of iron ore from the underground mines of the Lake Superior district is recognized, and appropriate measures for the protection of individuals subjected to these dusts have been applied in most mines for more than to years. Undoubtedly the principal concern has been demanded by dusts that contain as a constituent silicon dioxide in its free or uncombined state, the mineral we all know as quartz or chert. Investigators1 agree that among the factors that determine injurious exposure are: the composition of the dust, the concentration suspended at the breathing level and the duration of the exposure. We have on hand data that describe the experience of several underground iron-ore mines in the Upper Peninsula of Michigan and Wisconsin in regard to these factors, which in a large measure also reflect the general experience of the other mines of the area. We believe that a discussion of these data will indicate the effectiveness of the dust-control methods that have been considered good practice and that have been employed over a long period of time. A silicosis-prevention program was ini¬tiated by the iron mines of the Upper Peninsula in 1934 and records of mine dustiness are available from that date to the present time. For dust conditions previous to 1934 the results of observations conducted in other industries can be utilized together with local information pertaining to the development and initial application of rock-drilling equipment. DEVELOPMENT OF ROCK DRILLS The earliest percussion rock drill that later had any industrial use was patented in this country in 1851.2 This machine was later modified and improved in many respects but its piston-type design remained standard for drilling equipment in most mines for the next 63 years.3 The piston drill was operated either by air or by steam. It employed a solid drill steel, fastened directly to the reciprocating piston. The drilling was done dry except in holes directed at an angle below the horizontal. For these water was manually injected into the hole and the pumping action of the rod expelled the sludge. The dry cuttings from drillings inclined above the horizontal fell from the hole by their own weight. A development of major importance from the standpoint of dust generation and control occurred with the invention by J. G. Leyner of a hammer-type rock drill that embodied the essential feature of a drill steel with a hollow core. His patent was secured in 1897. The machine was a radical departure in drilling method. The steel was held loosely in a chuck while

Jan 1, 1947

By S. D. Warriner

ONE of the most, if not the most, difficult of the problems associated with the mining and distri-bution of anthracite is to get clearly into the minds of the consumers and of the editorial and reporto-rial writers in the public press the relationship between the cost of production, the price at the mine and the price that the consumer pays for his coal, delivered in the cellar. One sees in his morning paper a statement to the .effect that out of the $12 or $13 per ton that his dealer charges him, the miner gets only $2 or even less, and he asks himself or the paper asks for him, "Who gets the difference?" Not infrequently the amount that the miner gets is given for run-of-mine bituminous coal while the price quoted is that of prepared anthra-cite, egg, stove, or chestnut, as the case may be, delivered to the bin of the consumer; and the difference stands out with all the more prominence; and it sticks in his memory with a persistency which no manner of explanation is able to efface. A recently published official report for the year. 1921 of one of the State Bureaus stated the total production of anthracite in net tons of 2000 lb., and included not only the output recovered by washeries from old culm banks and by dredges from creek and river channels, but also the colliery consumption;, i.e., the coal used by the companies in the operation of the collieries. This resulted in a total production of 90,509,075 net tons for which the labor cost was stated at $283,961,300, or $3.13 per net ton, equivalent to $3.50 per gross ton. This figure of $3.50 was immediately seized upon as representing the labor cost per gross ton of fresh-mined commercial coal. It is almost a hopeless task to try to correct such errors of statement in the first place, and, in the second place, to correct the misapprehen-sions that arise from them. As a matter of fact, the report was misleading in regard to the labor cost of the entire production shipped, as the coal used for colliery consumption should be deducted. Correctly stated, the report should have shown, calculated in terms of gross tons of 2240 lb., the legal weight in Pennsylvania, as follows: Gross production 80,607,359 tons Less colliery fuel 8,757,137 tons Commercial production 71,850,222 tons Total labor bill $283,961,300.00 Labor cost per ton...: 8 3.952 The foregoing figures" of tonnage are obtained from the State Department of Mines. Even these include the quantity recovered from calm, banks by washeries, and consequently the $3.95 arrived at as the average labor cost is somewhat less than that actually involved in the mining of fresh-mined coal. Unfortunately, the figures of washery production and the cost of producing it are not available.

Jan 7, 1922

By M. J. Pool, P. J. Spencer

Using liquid tin as the solvent metal, the heats of solution of silver and gold in dilute Ag-Au-Sn alloys have been determined at 723°K by solution calorimetry. From the values obtained, the various enthalpy interaction coefficients in the liquid solutions have been calculated. The self-interaction coefficients Au and Ag values of 2413 and -6322 cal per g-atom in binary Au-Sn and Ag-Sn alloys, respectively. For an alloy containing 0.04 mole fraction Ag, Au is equal to 2443 cal per g-atom, and in an alloy containing 0.04 mole fraction Au, Ag has a value of —8883 cal per g-atom. Ag Au has been determined as —2650 cal per g-atom. ADDITION of a second solute to a dilute binary alloy may cause striking changes in the thermodynamic characteristics of the alloy. As yet, such influences cannot be reliably predicted and it is only by accumulation and critical assessment of data that definite patterns can be established. Solute interactions in dilute solutions were first treated mathematically by Wagner,1 who derived expressions for the Gibbs energy interaction coefficients. For an alloy consisting of. two dilute solutes, i and j, in a solvent s, these coefficients are given by: where yi is the activity coefficient of component i and x represents the mole fraction of solute or solvent. Other interaction coefficients have since been established in terms of the various partial molar thermodynamic properties of an alloy; the definitions of these are summarized in a recent paper by Lupis and Elliott.2 Mathematical and experimental studies of such coefficients have received much attention recently, providing a very useful theoretical background to an understanding of the nature of solute interactions taking place in dilute solutions. One of the definitions to arise from such work has been the extension of Wagner's treatment by Lupis and Elliott4 to include an enthalpy interaction coefficient, defined as. where ?Hi is the relative partial molar enthalpy of component i. ?Hi is related to the relative partial molar enthalpy at infinite dilution, ?Hio, and to the enthalpy interaction coefficients by the expression: Values of ?Hi can often be obtained with a high degree of accuracy by calorimetry, and, if the composition of the alloy is carefully controlled by making calculated solute additions throughout a series of measurements, application of Eq. [2] will enable reliable enthalpy interaction coefficients to be determined. Partial heats of solution must be investigated as a function of solute concentrations down to xi = 0, and, since ? values are obtained from the slope of mi plotted against x, great care must be taken to obtain precise data. The data for binary systems which are presently available often require extrapolation of ?Hi values from compositions as great as xi = 0.05 or even 0.1 to obtain limiting heats of solution. Reliable enthalpy interaction coefficients are thus impossible to obtain from published data for most of the alloy systems already investigated. The aim of the present work was to determine values of Au, Ag, Ag, and Au in the limit xsn -1 for the relatively simple Ag-Au-Sn system and from these results to gain some insight into the possible influence

Jan 1, 1969

By C. W. Hill

FLUXES, in general, may be classified according to their use as soldering, foundry or casting, and metallurgical and the chemistry of their action follows quite closely this division. The term foundry or casting fluxes, as used here, refers to the substances added to molten metals, preparatory to casting into molds or ingots, with the object of removing small amounts of impurities that have been introduced in handling or melting the metal, which was placed in a fair state of purity by the previous refining or metallurgical processes. The sources of these impurities are the oxide coating on the surface of the ingots, material extracted or mechanically removed from the crucible or furnace wall, and those formed by the reaction of the molten metal with the heating gases or furnace atmosphere. As commonly used, the term fluxes is applied to such materials as ammonium chloride, zinc chloride, salt, borax, and boric oxide. A second class of materials, usually called deoxidizers, includes substances, such as metallic phosphides, sub-oxides, carbides, and bo-rides, that reduce the metallic oxides present in the metal. A third class includes various compounds that are molten and non-volatile at the temperature of the molten metal and are used primarily to protect the surface of the metal from the action of the furnace or exterior atmosphere; to this class belong such mixtures as lime and fluorspar, lime and soda ash, various minerals, etc. In some cases these protective layers may act in a manner similar to the metallurgical fluxes and dissolve or unite with the oxide impurities of the metal, but as generally employed this latter action is negligible since there is little opportunity for the coverings to come into contact with more than the surface of the metal, and hence they are without action on the body of the molten mass. The use of fluxes is the result of factory experience and tradition, but an extensive study has not been made of the chemical, and physical actions that occur or the effects of these actions on the properties of the metal. It would seem necessary in any successful study of metal fluxes

Jan 10, 1920

By G. J. Harmsen

is shown that in any isotropic porous medium there is a fixed ratio, : , between the numerical valrles of the total pore perimeter per unit of cross-section and the specific surface area per unit of bulk volume. The consequences of this fact on the use of the concept "hydraulic radius" in the case of porous media are examined. The validity for consolidated sands of Kozeny's rela- tionship and of the formula derived by Carman for capillary rise are discussed. INTRODUCTION In the course of time several quantities and relations have been put forward in order to describe the texture of a porous medium and the flow of fluid through it. Among these relations the so-called Kozeny formula and a formula for capillary rise, for both of which Carman',' has given a derivation, will be specially considered here. These relations are of a semi-empirical nature and the derivations must be considered to have a heuristic significance only. For unconsolidated mate-

Jan 1, 1956

By A. E. Wraith, N. J. Petch, J. D. Embury, E. S. Wright

The two most important parameters controlling the fracture behavior of a solid are its intrinsic properties, e.g.,grain size, and the operative stress system. The latter may be modified in laminates by the presence of weak interfaces. This is studied in notch-impact tests on a mode1 system of mild steel laminates containing a variety of interfaces. The effect of these is evaluated in terms of the ductile/cleavage transition. Two laminate geometries are distinguished, here called "crack-arrester " and "crack-divider". In both, cleavage is inhibited. This arises because of relaxation of a state of triaxial tension. In the crack-avrester laminates, cleavage initiated at a notch is confined to the layer containing the notch. In crack-cliuider laminates, a thick specimen behaves as the Sum of a number of thinner ones. Additional benefit may derive from improved intritnsic propevties of- the lanlinate layers arising from greater deformation in their manufacture. It has long been recognized that the two most important parameters controlling the fracture behavior of solids are their intrinsic properties (e.g., grain size friction stress,3 distribution of second phase particles4) and the operative stress system under which fracture occurs. In solids that show a ductile/cleavage transition, cleavage is favored by the presence of a notch. This is because triaxial tensions, generated by the localized plastic constraint at the notch, are operative when fracture occurs.= Anything that suppresses these triaxial tensions will be unfavorable to cleavage. Such suppression may possibly occur in laminates containing weak interfaces and the purpose of the present paper is to explore this possibility. Two basic laminate geometries, here termed "crack-avrester" and "cvack-divider", are examined. They are illustrated in Fig. 1. With the crack-arrester laminates, there is the possibility that, when the fracture crack approaches the interface, this, if weak, may delaminate due to the tensile stress acting parallel to the plane of the crack.= If this happens, energy will be used in delamination, the crack will be completely blunted, and the triaxial tension associated with the crack will be relaxed. To fracture the second portion of the laminate, crack reinitiation will be necessary and, because of the relaxation of the triaxial tension, this reinitiation will occur under conditions of nearly uniaxial tension, which are unfavorable to cleavage. Thus there is the possibility of cleavage suppression in the second and subsequent subunits of a crack-arrester laminate. With the crack-divider geometry, there is again the possibility of delamination at the interfaces. This will divide the crack into a series of cracks propagating through the individual laminate subunits. If these are sufficiently thin, the triaxial tension will be relaxed towards a state of biaxial tension in each of them. Thus, with the crack-divider laminates, there is again the possibility of cleavage inhibition. In the present work, these possibilities are explored using a notched impact test on mild steel laminates bonded with soft solder, silver solder or copper. Even if delamination does not occur, it is still possible that cleavage may be inhibited in laminates. With the crack-arrester geometry, the cleavage crack in the first layer may be blunted and arrested by plastic deformation in the laminate bond, if this is ductile. Partial relaxation of the stress transmitted ahead of the crack into the second layer will then result and this will reduce the significance of this stress in the fracture of the second layer. With the crack-divider geometry, there cannot be much effect in the absence of delamination unless a large amount of energy is absorbed in rupturing the ductile material. EXPERIMENTAL DETAILS The composition of the mild steel (wt pet) was: 0.04 C, 0.29 Mn, 0.01 Si, 0.006 P, 0.008 S. "As-received" plate was annealed for 2 hr at 900°C and slowly cooled to give a grain size of 0.04 mm. The laminates were made by brazing or soldering together mild steel plates 8 by 3 in. by various thicknesses. These were obtained from the annealed plate by machining, so that the intrinsic material properties were .kept constant. Laminates containing two to six steel layers were studied using standard Charpy V-notch specimens cut from the bonded plates. Standard homogeneous specimens from the annealed plate and subsize ones from the laminate components

Jan 1, 1968

POSITIONS VACANT (Under this heading will he published notes sent to the Secretary of the Institute by members or other persons.) Mine and mill accountant for Mexico. Knowledge of Spanish desirable. Salary $150 to $200 gold per month with living quarters. No. 60. Master mechanic familiar with mine and cyanide plant machinery for Mexico, $200 gold per month and living quarters. No. 61. Electro-metallurgist for production of ferro-alloys, salary $200 to $400 depending on experience. No. 62. Man familiar with ferro-tungsten products of electric furnaces to draw up specifications for machine tools, to watch these tools in service and revise specifications accordingly. Salary $150 to $200 per month. No. 63. Man to take entire charge of the sales of a steel company making electric alloy steels in the shape of ingots, billets and bars. To the right man an opportunity will be offered for taking a leading position in the management of the company and of participating in its results. No. 68. Chemist wanted by Electric Steel concern in the East. Must be fully acquainted with alloy steel analysis. Send full particulars and references. No. 73.

Jan 3, 1916

IN presenting the medalist, Scott Turner, director of 'the Bureau of Mines, said in part: "Mr. Blaylock is a Canadian by birth, training, ex-perience and preference. He was born in the Province of Quebec, graduated from McGill University in 1899, and immediately thereafter became an employee of the company for which he is now general manager. His outstanding achievements have been the acquisition for his company of the Sullivan mine, and the organization of a metallurgical staff, which, under his leadership and personal direction, solved the difficult problems pre-sented by the extraordinarily complex lead-zinc-silver ores of the Sullivan mine. "By title only, I will suggest to you the high lights of his career: (1) Development, of the great mine at Kimberley, and a huge plant at Trail. (2) Discovery and application of a successful differential flotation process for Sullivan ore which had previously not re-sponded to any process. (3) Improvements in lead-blast-furnace and refinery practice. (4) Development of an electrolytic process for treatment of zinc concen-trates. (5) Development of an electric-furnace-smelt-ing-method for zinc-plant residues. Besides this, we recognize his great organizing ability, his gift of leader-ship, and the cooperative spirit he always displayed.

Jan 3, 1928

By D.E. Norquist, J. E. Moeller

The results of wear on marked balls, 4, 31/2, 3, and 2 in. diam are given. All balls were forged steel of practically the same chemical analysis and hardness. The results indicate that balls in a given mill for a given length of time will have equal diameter losses, regardless of size. IN order to determine the relative wear rates of several sizes of grinding balls, groups of 4, 31/2, 3, and 2 in. balls were marked individually and were charged, all at the same time, into each of two production mills grinding copper ore. The remainder, and vast majority, of the ball charge consisted of 2 in. diam, and smaller, white, cast iron balls. The original weight of each test ball was determined and recorded. Some of each group of test balls were recovered from the mills periodically, individually reidentified, and reweighed. The test balls were recovered during regular maintenance shutdowns, at approximately thirty-day intervals, so as to avoid disrupting operations. As soon as the weights had been recorded, each marked ball was recharged into the mill from which it had been taken. The test units were mechanically the same and theoretically received equal tonnages of feed. All of the test balls were of forged, alloy steel composition with only slight variations in chemistry that would not be expected to affect the physical properties. All were heat treated in the same manner to produce a high hardness, as equal as possible for all sizes and as uniform as possible from surface to center. Generally speaking, the hardness of the portion worn from the balls during the test ranged from 62 to 65 Rockwell C. The microstructure was, therefore, predominately martensitic. The method of marking the test balls consisted of forming a small, round hole, 3/16 or 1/4 in. diam, into the center of each test ball with a Thomas Metal Master disintegrator, which produces a hole in fully hardened steel without affecting the hardness of the surrounding metal. After the hole was formed, each test ball was individually weighed to the nearest gram and the weight was recorded. The balls were grouped as to size and each group assigned a code letter. The individual balls within the various groups were then assigned a number. A small copper disc was stenciled on one side with the group letter and on the other side with the individual ball number and dropped into the hole in the corresponding test ball. The hole was then plugged with a low-melting-point metal alloy, which could be melted out in boiling water. Half of the test balls in each group were charged into one mill and the other half into another mill, so that all test balls in each mill would be subject to, as nearly as possible, identical conditions for an equal length of time. Previous testing had indicated that the wear rates on balls of different sizes could not be accurately compared if weight loss, or percentage weight loss, were used as a basis for comparison. Fig. 1 is a graphic record of the average weight loss for each size group in mill No. 1, and it becomes readily apparent that the larger balls lose substantially more weight in a given length of time. Fig. 2 is a graphic record of the average percentage weight loss for each size group in mill No. 1, illustrating that the smaller balls definitely lose a greater percentage of their original weight in a given length of time. It was, therefore, decided to convert the actual weights of the test balls to diameter and follow the diameter losses periodically to see if the wear rates

Jan 1, 1951

By Hisaaki Takata, Nobukazu Nanko, Ichiro Izawa

INTRODUCTION Tochibora Mine belonged to Kamioka Mines, which are located in Gifu Prefecture, the cen- tral part of the Japanese main island, has occupied the important position as a main source of silver, lead and zinc for the smelter of Mitsui Mining & Smelting Co., Ltd.. The annual production of crude ore goes up to about 1 mil- lion metric tons with average grades of 20g/t Ag. 0.31% Pb, and 4.18% Zn. For the last 20 years, the mine has struggled with the depressed zinc metal price in yen and the escalation of operating cost, especially la- bour cost. For instance, prices of commodities and labour cost have been raised 8 times and 14 times respectively during this period. On the other hand, total operating cost has been con- trolled to increase 4.5 times. The productivity of the mine has been increased 4 times by means of improvement of individual works and adoption of perfectly mechanized trackless mining system to reform the mining method. Some improvements are presented as follows; Today, drifting is carried out by the tunnelling crew which consists of 9 operators equipped with 5 Mobile-Jumbos, 2 LHD's and a jeep for carrying and loading ex- plosives. In order to get the high utilization of these machines, each crew manages systematically and continuously adjacent 15 faces under the direction of a forceman. As a result, we have accomplished high advance of drift. Furthermore, raising is carried out by a set of the raise-borer in case of long raises like ore- passes and also by the stage blasting method in case of short ones like slot-cuts. Generally speaking, the sublevel stoping method is considered to be one of the best methods for various respects; high productivity, low cost, easy grade control and safety. This method has been introduced in Tochibora Mine since 1950 and is now widely applied to lodes that could not be mined before by this method be- cause of the shapes of them and the conditions of their country rock. Now we are convinced to be nearly accomplished for our method. This report presents some examples of the standard sublevel stoping method and its variations which are practised at the mine. STANDARD SUBLEVEL STOPING Up to 1979 since 1950 when this method was first applied at the mine, 17 million tons of crude ore have been produced by this method, which amounted to 55% of the total production of the mine. During three decades, the productivity has been increased considerably by means of trackless mining system using large-sized machines and etc.. In other words, mining efficiency has been increased from 20 tons/man.shift to 100 tons/man.shift at large stopes. Formerly undercutting was carried out above the drawing drift by jack-legs and small scrapers and then several cone-shaped glory holes were made by stopers between the undercut level and the draw level. But now, a undercut drift is driven in parallel with the haulage drift at the same level of it. And at the undercut drift, drilling upward is carried out by a large-sized machines like the fan drill in order to extract the broken ore directly at the drawing drift connected with the haulage drift. Simplification of the work as well as utilization of large- sized machines enables to get higher productivity. But if the deposit is comparatively wider and the broken ore around the middle part of the deposit cannot be extracted, several drawing drifts are driven to the deposit and in order to minimize dead ore at the final stage of operation, the

Jan 1, 1981

By D. E. Norquist, J. E. Moeller

The results of wear on marked balls, 4, 31/2, 3, and 2 in. diam are given. All balls were forged steel of practically the same chemical analysis and hardness. The results indicate that balls in a given mill for a given length of time will have equal diameter losses, regardless of size. IN order to determine the relative wear rates of several sizes of grinding balls, groups of 4, 31/2, 3, and 2 in. balls were marked individually and were charged, all at the same time, into each of two production mills grinding copper ore. The remainder, and vast majority, of the ball charge consisted of 2 in. diam, and smaller, white, cast iron balls. The original weight of each test ball was determined and recorded. Some of each group of test balls were recovered from the mills periodically, individually reidentified, and reweighed. The test balls were recovered during regular maintenance shutdowns, at approximately thirty-day intervals, so as to avoid disrupting operations. As soon as the weights had been recorded, each marked ball was recharged into the mill from which it had been taken. The test units were mechanically the same and theoretically received equal tonnages of feed. All of the test balls were of forged, alloy steel composition with only slight variations in chemistry that would not be expected to affect the physical properties. All were heat treated in the same manner to produce a high hardness, as equal as possible for all sizes and as uniform as possible from surface to center. Generally speaking, the hardness of the portion worn from the balls during the test ranged from 62 to 65 Rockwell C. The microstructure was, therefore, predominately martensitic. The method of marking the test balls consisted of forming a small, round hole, 3/16 or 1/4 in. diam, into the center of each test ball with a Thomas Metal Master disintegrator, which produces a hole in fully hardened steel without affecting the hardness of the surrounding metal. After the hole was formed, each test ball was individually weighed to the nearest gram and the weight was recorded. The balls were grouped as to size and each group assigned a code letter. The individual balls within the various groups were then assigned a number. A small copper disc was stenciled on one side with the group letter and on the other side with the individual ball number and dropped into the hole in the corresponding test ball. The hole was then plugged with a low-melting-point metal alloy, which could be melted out in boiling water. Half of the test balls in each group were charged into one mill and the other half into another mill, so that all test balls in each mill would be subject to, as nearly as possible, identical conditions for an equal length of time. Previous testing had indicated that the wear rates on balls of different sizes could not be accurately compared if weight loss, or percentage weight loss, were used as a basis for comparison. Fig. 1 is a graphic record of the average weight loss for each size group in mill No. 1, and it becomes readily apparent that the larger balls lose substantially more weight in a given length of time. Fig. 2 is a graphic record of the average percentage weight loss for each size group in mill No. 1, illustrating that the smaller balls definitely lose a greater percentage of their original weight in a given length of time. It was, therefore, decided to convert the actual weights of the test balls to diameter and follow the diameter losses periodically to see if the wear rates

Jan 1, 1951

By R. F. Mehl, C. E. Birchenall

R. E. Hoffman and D. Turnbull—The authors have presented evidence which they have interpreted as indicating that the rate of self diffusion is not intrinsically more rapid at grain boundaries than within the grain. Grain-size effects which apparently exist are attributed rather to impurities concentrated at the grain boundaries. In view of our own experiments and the existing evidence, we believe that the support for this hypothesis is not convincing. We have in progress an investigation in which the rate of self diffusion of silver is being measured over an extended temperature range in both single-crystal and polycrystalline specimens. The results of the single-crystal experiments and some preliminary data on fine-grained polycrystalline specimens have already been reported:' and it is anticipated that a complete report will be published in the near future. The self-diffusion coefficient of large-grained polycrystalline silver (1 grain per sq mm) has previously been measured by Johnson" between 730" and 940°C. The diffusion coefficients which we have measured in single crystals (210 plane normal to diffusion direction) agree within experimental error with values calculated from an extrapolation of Johnson's curve down to temperatures as low as 500 °C. However, it has been demonstrated that the overall self-diffusion rate in fine-grained polycrystalline specimens (initial grain size of 0.003 cm) becomes measurably larger than the overall rate in a single crystal at a temperature of 600°C, and the discrepancy between the two rates becomes greater as the temperature is further decreased. In fact, it has been possible to obtain satisfactory penetration curves for polycrystalline specimens using the sectioning technique at temperatures as low as 400°C. At this temperature, the penetration is 50 to 100 times greater in the polycrystalline specimens than in a single crystal. Fisher" has developed an analysis whereby the ratio of the rate of the unit diffusion process at the grain boundary to the corresponding rate within the grain can be calculated from the penetration curves and an assumption as to the width of a grain boundary. This analysis applied to our data indicates that the unit process at the grain boundary is faster by a factor of 10' at 475°C when the grain boundary width is taken to be 5. The silver used in most of these experiments was obtained from the Handy and Harmon Co. and listed as 99.97 pct pure. Preliminary experiments on 99.999 pct silver from the Jarrel-Asch Co. indicate a grain-size effect of the same order of magnitude as in the less pure silver. Nominally, these purities are as good, at least, as that of the carbonyl iron used by the authors, but of course if an impurity effect does exist its magnitude might be very dependent upon the nature of both major and minor constituents. The authors have cited the work of other investigators who have found no grain-size effects. Neither Steigman, Shockley and Nix" nor Maier and Nelson8 were able to correlate self-diffusion coefficients of copper with grain size. However, all their measurements were performed at or above 750°C; and on the basis of our work with silver, no grain-size effect would be expected at temperatures above about 0.7 of the absolute melting temperature unless the grain size were exceedingly small. Likewise, in the investigation of the self diffusion of lead by Seith and Keil,14 the lowest temperature at which the diffusion coefficient was measured in polycrystalline specimens was 207°C, which is still sufficiently high so that the lack of a grain-size effect is not surprising. Finally, in those experiments on iron from which they concluded that there was no grain size effect, Drs. Birchenall and Mehl seem to have no information as to the actual grain sizes immediately prior to and following the diffusion anneal. Without this information, we believe that their own experiments offer little support for their hypothesis. F. S. Buffington, I. D. Bakalar, and M. Cohen—The results given in this paper agree in order of magnitude with those tentatively reported by us.27 However, significant differences exist in the two sets of data, and it may be well to make an explicit comparison. The diffusion studies at M.I.T. were conducted on somewhat higher purity iron (99.98 pct Fe) than the grades used by the authors, but this is undoubtedly not the answer. Fig. 4 shows the diffusion results of both laboratories for the gamma phase, omitting the authors' data on the commercial steels, while fig. 5 presents a similar comparison for the alpha phase. The divergence is much more marked in the latter case than in the former. In connection with the M.I.T. determinations, all of the runs in the gamma range and those above 800 °C in the alpha range were conducted with specimens having a relatively thick (0.002 cm) electrodeposit of radioactive iron. This practice minimizes any possible error due to extraneous diffusion that may occur during the heating to and cooling from the operating temperature. An exact solution of Fick's law for these boundary conditions was used in calculating the diffusion coefficients. At a later time, three runs were made below 800°C, using very thin electrodeposits similar to those of the authors, and the points fell considerably below the values expected from the extrapolation of the results based on the specimens with the thick deposits (compare dash-dot line in fig. 5). However, in the runs with the thin deposits, deviations of 100 pct were found between the individual specimens, whereas the maximum deviation with the thick deposits was less than 25 pct. Accordingly, it is not known at the moment whether the M.I.T. points below 800 °C should be given as much weight as those above 800°C. If this were done, the frequency factor would be of the order of 400 cm2 per sec, which is quite high. In other metals, the frequency factor for self diffusion lies between about 0.1 and 10 cm2 per sec. As the points below

Jan 1, 1951

By R. Hogg

Sintering is a process in which the pore structure of an agglomerated powder is modified. For many applications, it is desirable to be able to control the pore structure in order to optimize such properties as mechanical strength and reactivity (availability and accessibility of solid surfaces to reactant gases, for example). In general, these properties will be determined by the total porosity and the pore size distribution of the agglomerate. In the work described in this paper, a physically realistic mathematical model for the sintering process has been developed. The model is based on the usual assumption that the pore system in an agglomerate can be treated as a network of interconnecting, cylindrical pores of continuously varying size. It is further assumed that the total pore length is conserved during the course of the sintering process. Using standard expressions for the rate of shrinkage of an individual pore, the model can be used to predict the changes in total porosity and pore size distribution as sintering proceeds. Experimental pore size distributions, obtained by mercury intrusion porosimetry, for the sintering of oxides appear to be in general agreement with the predictions of the model.

Jan 1, 1977

By Charles Macdonald

Wrought-iron eye-bars for bridges and roofs, designed upon what is known as the pin connection system, have been successfully manufactured in this country for some years. The most approved methods employed have been the hydraulic upsetting process, and by die-forging under a hammer. In upset eyes the fibre of the iron is necessarily much distorted ; and in bars of considerable size, this disturbance seriously impairs the strength of the connection. Mr. Collingwood, in discussing my paper on "Proportions of the Heads of Eye-bars," read before the American Society of Civil Engineers, Jan. 21st, 1874, submitted sketches of certain bars intended for the anchorage of the East River bridge, which had been exposed to the action of sulphuric acid a sufficient length of time to thoroughly cleanse the surfaces, and expose the condition of the fibre. These sketches are here reproduced for the purpose of illustrating the effect of this method of manufacture. In Figs. 1, 2, and 3, it will be observed that the fibres at A, and E, have been folded back upon: themselves and otherwise twisted and disturbed, in a manner not calculated to inspire confidence in the integrity of the connection. To compensate for this defect, a large excess of metal is thrown into the eye; but a method of manufacture which involves such a necessity, even although it be the one giving the best results thus far, all things being considered, is far from being perfect.

Jan 1, 1879

By S. T. Wlodek, John Wulff

Dense, adherent, and ductile coatings of chromium can be applied to Molybdenum by selectively freezing out the chromium solute from a supersaturated copper-or tin-rich liquid alloy. The successful exploitation of this technique depends on the control of variables common to any freezing process. Under the most favorable combimtion of supersaturation, nucleation, and mass-transfey, coatings as thick as 0.004 in. can be obtained. In exploratory experiments it was found that a coating of chromium may be applied to molybdenum by immersing the molybdenum body in a supercooled copper-chromium or tin-chromium melt and selectively freezing out the contained chromium onto the molybdenum body. Supercooling of the liquid metal plating bath can be achieved by cooling, or the maintenance of thermal or concentration gradients in the melt. The technique is similar to conventional dip-plating procedures where the sample is immersed in a molten bath of the pure plating medium, except that in these studies, the plating medium consisted of a dilute solution of chromium in liquid tin or copper. The success of this method requires that the base material to be plated be isomorphous with the coating and insoluble in the plating bath. In addition, the plating bath, while having an appreciable solubility for the coating component, must have negligible solid solubility in the solid plate. Such relationships exist in the molybdenum-chromium, chromium-tin, and chromium-copper systems. Liquid tin and copper are satisfactory as the plating solvent, because although chromium is appreciably soluble in both of these liquid metals, solid inter-solubility is almost negligible, and no compound formation occurs. The same relationships exist between molybdenum and copper or tin except that the solubility of molybdenum in liquid tin or copper is insignificant. It is this combination of solubility relationships, as well as the completely isomorphous nature of the molybdenum-chromium system, which allows the formation of a chromium coating on molybdenum by the selective freezing mechanism. Since the main deterrent to the widespread use of molybdenum for high-temperature structural applications is its poor oxidation resistance, the formation of a dense and ductile chromium coating on molybdenum may well be of commercial significance. In addition, the basic principle of coating formation by selective freezing should be applicable to other metallic or even nonmetallic systems. The precipitation of the solute from a liquid metallic solution, for the formation of a protective coating, has not been previously considered in the literature. Frequently, however, a similar phenomenon has been described in the corrosion of metallic heat exchangers by liquid metals1 wherein temperature gradients in the liquid metal heat-transfer medium result in the formation of layers or deposits of metallic residue2 in the cooler portions of the system. Impregnating techniques using liquid baths have been tried as a plating process. Semchyshen and Barr, for instance, have successfully diffused chromium into molybdenum by immersing the molybdenum specimen in a liquid tin-chromium or copper-chromium alloy. This simple holding technique fails to produce a layer of pure chromium and requires prolonged immersion to allow for the diffusion-activated enrichment of the molybdenum. During the course of this investigation, most of the phase relations published for the copper-chromium4 and molybdenum-chromium4 systems, were found reliable. It was necessary, however, to investigate the solubilities of chromium and molybdenum in liquid tin as well as the solubility of molybdenum in liquid copper. These determinations were made by a simpie weight loss technique on materials of the same purity as used in the plating studies. The results of these solubility studies are given in Tables I and I1 and indicate that although the solubility of molybdenum in liquid

Jan 1, 1961

By L. S. Kolbert, C. H. Baab

Introduction In June 1981, the Electric Power Research Institute (EPRI) issued a report titled "Control Systems in Coal Preparation Plants" to determine the status of instrumentation and automation systems in the coal preparation industry. EPRI concluded that the control systems presently being employed in the US coal preparation plants "are relatively primitive compared with those in other mineral processing industries; and analytical instrumentation is virtually non-existent." With tighter specifications on the quality of steam coal, interest is increasing in more sophisticated control systems. One type of device that is currently assisting the preparation plant operator optimize and control the coal cleaning process is nuclear based instrumentation. Nuclear instrumentation includes methods for measuring weight, level, and density as well as for performing elemental analysis on coal. Major Features and Benefits of Nuclear Measurement and Analysis Systems There are many features and benefits of nuclear instrumentation. The major benefit is that nuclear instrumentation is non-contacting. The measuring heads do not contact the process materials. This enables measurements of corrosive, abrasive, or high temperature materials with no loss of performance, reliability, or useful life. Also, there are no moving parts to wear or be fouled by dust or corrosion. Installation is simple and inexpensive. For instance, density gauges are simply clamped onto existing pipe runs. Weigh scales and level switches usually require nothing more than welding brackets to customers' equipment. Once installed, routine maintenance can be performed without interrupting the process. Thus, nuclear instrumentation is highly reliable and easily installed and maintained. The operating life of the instrumentation is virtually unlimited. Principle of Operation Nuclear belt weigh scales, point level switches and density gauges all operate on the gamma-ray transmission principle. A beam of gamma radiation from a radioisotopic source is projected through the process material either on a belt, in a vessel, or in a pipe. Opposite the source is a radiation detector whose electrical output is proportional to the intensity of the radiation it absorbs. Gamma radiation is part of the same electromagnetic spectrum that encompasses light, infrared, and x-rays. In many respects, gamma radiation behaves much like light. However, it is far more penetrating and can be transmitted through considerable thickness of materials that are opaque to visible light, e.g., several inches of steel. Types of Nuclear Instrumentation in Coal Preparation Plants There are four types of nuclear instrumentation presently being employed in coal preparation plants. They are: •Belt Weigh Scales •Point Level Switches •Density Gauges •Coal Analysis Equipment. Each type of instrumentation has a number of potential applications within the plant. What follows is a discussion of three different types of devices and their applications. Weigh Scales In coal preparation plants, weighing instruments are used for measuring material throughput, totalized weight, and batch weight on conveyor belts and other transport equipment. They are also used on drag chain conveyors, vibrating conveyors, and screw conveyors. Weighing is important for receiving the run-of-mine coal, controlling mass flow, scheduling production, and loading and shipping the washed coal to market. There are two basic types of weighing instruments: gravimetric and nongravimetric devices. Gravimetric devices are units which measure the pull of gravity on the material being weighed. Nongravimetric devices are not dependent on the force of gravity on the sample. Nuclear weigh scales are the most prevalent nongravimetric devices used. The nuclear weigh scale works on the same basic principle as other nuclear gauges. The greater the quantity of material placed in the path of a radiation beam, the more radiation will be absorbed by the material. In weigh scales, Cesium 137 source is generally placed above the moving stream of material so that the radiation is directed through the material. An ion chamber radiation detector is on the opposite side of the material. The radiation passing to the detector generates a signal which is inversely proportional to the weight of the material present. This signal is amplified and integrated with the speed of the process flow yielding the weight. The weight is indicated and can be totalized for throughput, batch control, etc. In conclusion, radiation weighing methods have become well established in the coal preparation industry throughout the world. As mentioned before, this device finds its most

Jan 1, 1985

By G. W. Healy, L. F. Sander

was investigated by reacting thin discs of calcium oxide with gas mixtures of CO2, CO, and Son. Its value was 19,300 * 300 cal independent of temperature in this range. No solid solubility of sulfur in calcium oxide was detected within the limits of the experimental method and it is estimated to be below 0.025 pct by weight. The importance of lime in desulfurization is well-established but complete information on the pure phase equilibrium: CaO + 1/2 s2 = CaS + +02 [11 is not yet available. The goal of this work was to evaluate solid solubility of CaS in CaO and to determine the free-energy change associated with Reaction [I] at temperatures of 1400" to 1650°C. The equilibrium constant for Reaction [1] can be written: It is convenient to rewrite Eq. [2] in the form: where A = {Ps /PqJ1'2 has been referred to' as the "sulfurizing power' of a gas mixture. In this work, thin discs of CaO were suspended in a vertical tube furnace and exposed to CO + CO2 + SOz gas mixtures having known values of A. The samples were then analyzed for sulfur. As expected, X-ray diffraction confirmed that CaS was the only sulfur-bearing phase formed at the relatively low oxygen pressures used. EXPERIMENTAL PROCEDURE Reagent-grade CaCO3 was pressed in a 3/8-in.-diam pill die and prefired in air to produce CaO discs weighing between 0.004 and 0.01 g. Several discs were used to provide a suitable weight for chemical analysis while maintaining a large surface area to react with gas mixtures. These were placed in a platinum mesh basket and suspended in the gas stream in the hot zone of a vertical tube furnace. Desired gas mixtures were prepared from cp grade CO and CO2 and anhydrous grade SO2. The method of soap bubble displacement was used to calibrate capillary flow meters. While this gave excellent results with CO and Con, some problems with bubble insta- bility and soap film "drag" arose with the use of SO2 at low flow rates. Hence, frequent sampling and analysis of gas mixtures was carried out to insure proper control of the ingoing SOZ. The furnace used for gas:solid equilibration was a vertical mullite tube externally wound with 60 pct Pt-40 pct Rh wire having a diameter of 0.028 in. An inner tube of $ in. ID served as the reaction chamber having Pyrex ground joints sealed to the mullite to provide gas-tight connections at top and bottom. A Pt-Pt 10 pct Rh thermocouple was inserted into a protection tube adjacent to the sample basket to measure sample temperature during a run. Constant-temperature control to 2C was observed at any desired set point within the range of this investigation. This was accomplished by a control thermocouple imbedded in the furnace windings which served to actuate an electronic controller wired for high-low operation. The sulfur analyses of the solid samples were carried out using a stoichiometric combustion technique based on the method of Fincham and Richardson. Some analyses were done using a modified evolution method3 but these were used primarily to check the results of the combustion method. The results were in good agreement but the combustion technique of-ferred an advantage in economy of time and material. CALCULATION OF GAS EQUILIBRIA Heating a given mixture of CO + CO + SO2 to high temperatures gives rise to a large number of product species. The details of calculating the partial pressures of these products of interaction and dissociation can be found in several references4,5 and need not be repeated here. The thermodynamic data selected for the major species in the gas mixtures are shown in Table I. Equilibrium constants from these reactions were combined with oxygen, carbon, and sulfur balances and a computer program written to facilitate the calculations. Some early difficulties in reproducing experimental results were finally traced to the effect of atmospheric pressure changes. No reference to consideration of this question had been found in the

Jan 1, 1969

By Albert Hahn

ABOUT three years ago John Hays Hammond took over the control of the Eureka Metallurgical Co., at Salt Lake City, Utah. Funds were advanced for investigating the process invented by R. V. Smith, for concentrating oxidized lead ores. A great deal of laboratory work was done and a reagent was developed by C. M. Nokes of Salt Lake, which was found to be particularly suitable for oxidized copper ores, refractory manganese silver ores, and oxidized gold ores. After sufficient laboratory work had been performed to demonstrate the fundamental principles of these inventions, a test mill, having a capacity, of 20 tons of ore daily, was erected at Murray, Utah. In this test mill approximately 1200 tons of various types of ores have been concentrated by means of the above-mentioned process and reagents, and this large-scale work as well as laboratory work is being continued at present. Details of these mill tests are given in Table 1, and a representative group of laboratory tests are shown in Table 2. The process requires no special machinery or appara-tus. Its success depends entirely, on the method of preparing the pulp and its subsequent flotation in clean water, as outlined in the Smith patent. The reasons for the success of the Nokes special reagent are not definitely known. The outstanding feature, which seems to be of extreme importance, is that when the concentrates are examined under the microscope, each particle of mineral appears to have a particle of paraffin, attached to it. This paraffin-mineral combination seems to be particularly suitable for flotation and it follows that any coating oils readily attach themselves to the paraffin-mineral combination.

Jan 9, 1923

By J. D. Audley Smith

Heine's " blue test" for copper, as described by the authorities generally, calls for a set of standard colors; and there has been some discussion concerning the relative superiority, for this purpose, of sulphate and nitrate solutions. The whole matter evidently hinges on the preservation of the standard colors in well-stoppered bottles. The apparatus described by Q. L. Heath* cannot be much improved, when very accurate readings are required. The following method was devised for the purpose of doing away with preserved sets of standard colors, by making a fresh standard for each batch of assays. The solutions are prepared in the usual way, each ammoniacal solution being filtered into its separate bottle, and then filled up to the containing-mark and thoroughly mixed. A similar color-bottle, with an " S " etched upon it, to distinguish it from the other bottles, which are numbered to correspond with the samples, is kept to run the standard. About 150 C.C. of water is put into it, and then the amount of acid (sulphuric or nitric) present in each determination, followed by 30 C.C. of ammonia (s. g. 0.90), which should make the mixture strongly ammoniacal. The liquid is now made up almost to the containing-mark (200 c.c.), say within 1 C.C. The lowest assay is selected first and placed alongside the standard, which, at the beginning, contains no copper. A copper-solution is then dropped into the standard from a burette ; and after each addition the bottle is well shaken and compared with the assay-sample. This operation is repeated until the two shades match exactly, when the burette is read and the assay-result is calculated. The assay next in order of depth of color is now taken and treated in the same way, and so on, until the batch is completed. It

Jan 1, 1901