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By B. B. Gottsberger

IT SEEMS strange that so many years after the pas¬sage of the first acts requiring registration or licensing of engineers, so few members of the mining branch of the profession are aware of what has taken place. Laws regulating the practice of engineering have now been passed in 19 states of this country, and in many others, such legislation has been proposed. That such laws are necessary for the protection of life, health and property, as is stated in most of them, is doubted by many. That the standing of the engineering profession will be improved by them is certainlyy open to question. That such laws are a welcome addition to the regulation of business by governmental agency, is not likely to be a general opinion. That once more, and in spite of appeals to engineers to interest themselves in affairs of government, an active minority has succeeded in imposing its will on a passive majority, seems but too true. However, the laws are there and information regarding them should be of use and interest to those most directly concerned.

Jan 1, 1921

By William S. McAleer

Two major trends in the coal industry today focus attention on the need for heat-drying equipment of a simpler, more flexible and less expensive type than has been considered standard equipment for drying coal. These trends are: I. The tremendous growth in the use of mechanical equipment to load coal both in underground mining and strip mining, which necessitates efficient mechanical cleaning of the fine as well as the coarse coal and thus means a growing tonnage of washed fine coal? with its attendant moisture problems. 2. The growing consumer demand for the smaller sizes for use as domestic stoker coal, with increasingly rigid specifications covering size, ash, and moisture analyses. ACCEPTED PRACTICE In the past, when, for example, 4 in. too coal has been washed, there was little trouble in producing a washed product 4 to o with a satisfactory resultant moisture. The 4 to 3/8-in. could be taken from dewatering screens at a surface moisture ranging from 2.5 to 3.5 per cent, depending upon the size consist. The 3/8-in. to o coal was generally dewatered satisfactorily with centrifugal driers, to final surface moistures from 5.5 to 8 per cent, again depending upon size consist. The centrifugal type of drier has definitely proved itself to be the most economical type of machine to handle coal approximately 3/8 in. to o as recovered by dewatering screens or drainage elevators with 15 to 25 per cent surface moisture and turn out a product with moistures as stated above. Sometimes such moistures (5.5 to 8.0 per cent) in the product are satisfactory. If not, an ultimate moisture, lower than 3 per cent, is seldom required and the centrifugal drier removes 70 to 85 per cent of the total moisture to be removed. In certain districts there has been a growing tendency to circumvent the use of centrifugals or other types of driers, by dewatering over screens of the wedge wire type and wasting all the finer sizes. Sometimes all the minus 28-mesh has been wasted and often the size wasted has been as large as 1/8-in. By wasting these finer sizes, a lower moisture can be obtained in the finished product, but taking into account the cost of the tonnage so lost, plus the added cost of ultimately disposing of this wasted tonnage, often the cost will be greater than the cost of heat-drying in a modern plant, which makes possible the recovery of all the tonnage. Moreover, with the moisture standards for domestic stoker equipment even these wasteful approaches to its solution do not solve the problem and there remains only the alternative of heat-drying. HEAT-DRYING Until about four years ago, heat-drying of coal at the preparation plants was confined chiefly to drying the fine sizes (approximately 3/8 in. to 0). Occasional

Jan 1, 1941

By Hugh Marriott

MANGANESE and its ores have been recently dealt with in comprehensive papers to the Iron and Steel Institute by Sir Robert Hadfield, and in a series of papers read before the A. I. M. E. at the Cleveland meeting. I have taken the opportunity of adding this mineral, to my records of the "World's Mineral Production and Consumption," as recently pub-lished in the Economist ?. The tables bring out the following significant facts: The world's production has increased by 10 per cent since 1913. India's production has increased by 3 per cent since 1.913, the supply from this source now being 31.1 per cent of the world's output. Russia is now producing at about half its pre-war rate and yields 19 per cent of the world's production. West Africa supplies 12.5 per cent of the world's total output. Brazil's present output is 2.5 times that of the pre-war period, being 11.3 per cent of the world's output.

Jan 9, 1927

By J. W. Edington

THE Hopkinson bar has become a popular technique for the measurement of the mechanical properties of materials deformed at high strain rate. Maximum use of the equipment is made in the arrangement first used by Kolskyl in which a short compression specimen is sandwiched between two pressure bars and is loaded by a single pulse travelling through the system. The pressure bars are used both to apply the load to the specimen and as transducers to obtain continuous strain-time histories of three pulses, incident on, reflected, and transmitted by the specimen. The data measured by the pressure bars can be analyzed in terms of the stress/strain behavior of the specimen.2'3 However, one of the assumptions of the analysis of the observed pulses is that the total stress and total strain do not vary significantly from point to point within the specimen at any given instant during the deformation process. Although this assumption is generally justified for very short disc-like specimens2 the situation is uncertain for larger specimens. For example, at small plastic strains (-0.01) Hauser et al.2 have some evidence of small flucations in the total stress within the crystal during deformation, even in relatively short aluminum specimens. In addition, Karnes4 has shown that the plastic strain, and by inference strain rate, is different at each end of a compression specimen tested in a Hopkinson bar, although the length of the specimen was not specified. Recently, the mechanical properties and the dislocation substructure have been investigated in single crystals of columbium5 (length 0.25 in., diam 0.19 in.), and copper6 (length 0.5 in., diam 0.5 in.) deformed at high strain rates. As part of this research program the assumption that the plastic strain is constant throughout the specimen has been checked by measuring the total dislocation density as a function of position in the specimen. Compression specimens of the same orientations and dimensions were tested as described previously5,6 sing a split Hopkinson bar. Since any discontinuity in strain distribution is most likely to arise during the initial stages of deformation the investigation was performed on specimens deformed to plastic shear strains of 0.054 (copper) at a strain rate 1.2 x l03 sec-1, and 0.06 (columbium) at a strain rate 1.5 X l03 sec-1. The orientation of the single crystals is shown in Figs. 1 and 3. Thin foils were taken parallel to the most highly stressed slip plane, i.e., (111) in copper and (011) in columbium, using conventional disc techniques. The dislocation densities were measured using first order reflections with compensation for invisible dislocations.5'6 In the copper single crystals the discs were randomly distributed throughout the cross section of the specimen. However, the dislocation density obtained from each disc was plotted vs the disc positions relative to the ends of the specimen. The results for the copper specimens are shown in Fig. 1. Clearly the dislocation density is constant throughout the main portion of specimen within the experimental error. The error bars on the dislocation densities correspond to a shear strain variation of 0.015 on the basis of previous measurements% ± of the rate of increase of dislocation density with strain in copper single crystals of the same geometry. Thus within this experimental error the plastic strain can be concluded to be constant within the specimen and the assumptions used in the analysis of the stress/time curves are therefore reasonably valid. The higher measured dslocation density near the impact end and the lower dislocation density at the bar end of the copper specimen is in agreement with the results of Karnes4 who showed that this strain/time curve rose to a maximum more rapidly at the impact end compared with the bar end. Hauser et al.2 have also pointed out that at small plastic strains (-0.01) the strain at the impact end of the specimen may be greater than that at the bar end. Thin foils taken from different points within the columbium single crystals demonstrated that the dislocation density could vary significantly within the specimen, see Fig. 2. Large areas of some thin foils up to 30 µ sq contained very few dislocations, see Fig. 2(a). However, in other parts of the compression specimen dislocation configurations like those shown in Fig. 2(b) existed over large areas (-30 µ sq). As a result, when the average dislocation density in a thin foil is plotted as a function of the position of the thin foil relative to the ends of the specimen, considerable scatter is observed, see Fig. 3. In this material then, the local dislocation density, and consequently the

Jan 1, 1970

By N. Van Wingen

EFFective sand permeabilities can be ascertained from core analysis if the laboratory data are compensated to allow for the presence of connate or residual water. Such adjustments can be made by applying empirically derived correction factors. Effective sand permeabilities can be estimated also from analytical studies of field depth-pressure measurements. In one method the time rate of pressure build-up after a well is shut in from steady-state flow is required, and in the other the direct measurement of a steady-state rate of flow and the corresponding equilibrium pressure. The results of the investigation discussed in this paper demonstrate that these methods for determining mean effective sand permeabilities can be correlated. Introduction In recent years many investigators have demonstrated the practical value of an accurate knowledge of the physical and dynamical characteristics of reservoir sands containing oil and gas. Of these factors, which serve to completely define a porous medium as a container or carrier of fluids, the dynamical characteristics are of particular interest to the production engineer because they afford a direct measure of the rate at which the oil or gas content of a reservoir will become available. The constant that defines dynamically the porous medium as the carrier of a homogeneous fluid in viscous motion is the "permeability." Coke-analysis Method The generally accepted practice in com-piling permeability data has been to analyze in the laboratory core samples taken in the course of drilling through the productive horizon, and to obtain the mean measure of the productive ability or permeability for the interval open to production by graphical integration of the permeability profile. An average permeability ascertained in this manner, however, is useful to the production engineer only for comparisons of wells drilled in the same horizon, as permeabilities determined in the laboratory are measures of only the ability of a porous medium to transmit a single fluid. As a result of many investigations, it has been shown, however, that under actual reservoir conditions water is coexistent with the petroleum and that its presence greatly reduces the permeability of the sand to oil. Empirical relationships showing the effect of the presence of any immiscible fluid A (such as water) in a porous medium on the permeability of the medium to a fluid B (such as oil) have been established separately by Hassler,1 Dunlap,2 and the writer (Fig. I),3 so that the average specific or dry-core permeability as determined in the laboratory can be corrected to give the true measure of the ability of the horizon penetrated to transmit oil, provided the percentage of pore space occupied by connate water is known accurately.

Jan 1, 1942

By N. Van Wingen

EFFective sand permeabilities can be ascertained from core analysis if the laboratory data are compensated to allow for the presence of connate or residual water. Such adjustments can be made by applying empirically derived correction factors. Effective sand permeabilities can be estimated also from analytical studies of field depth-pressure measurements. In one method the time rate of pressure build-up after a well is shut in from steady-state flow is required, and in the other the direct measurement of a steady-state rate of flow and the corresponding equilibrium pressure. The results of the investigation discussed in this paper demonstrate that these methods for determining mean effective sand permeabilities can be correlated. Introduction In recent years many investigators have demonstrated the practical value of an accurate knowledge of the physical and dynamical characteristics of reservoir sands containing oil and gas. Of these factors, which serve to completely define a porous medium as a container or carrier of fluids, the dynamical characteristics are of particular interest to the production engineer because they afford a direct measure of the rate at which the oil or gas content of a reservoir will become available. The constant that defines dynamically the porous medium as the carrier of a homogeneous fluid in viscous motion is the "permeability." Coke-analysis Method The generally accepted practice in com-piling permeability data has been to analyze in the laboratory core samples taken in the course of drilling through the productive horizon, and to obtain the mean measure of the productive ability or permeability for the interval open to production by graphical integration of the permeability profile. An average permeability ascertained in this manner, however, is useful to the production engineer only for comparisons of wells drilled in the same horizon, as permeabilities determined in the laboratory are measures of only the ability of a porous medium to transmit a single fluid. As a result of many investigations, it has been shown, however, that under actual reservoir conditions water is coexistent with the petroleum and that its presence greatly reduces the permeability of the sand to oil. Empirical relationships showing the effect of the presence of any immiscible fluid A (such as water) in a porous medium on the permeability of the medium to a fluid B (such as oil) have been established separately by Hassler,1 Dunlap,2 and the writer (Fig. I),3 so that the average specific or dry-core permeability as determined in the laboratory can be corrected to give the true measure of the ability of the horizon penetrated to transmit oil, provided the percentage of pore space occupied by connate water is known accurately.

Jan 1, 1942

By C. WIFLARD HAYES

(New Haven Meeting, February, 1909.) I DESIRE to make it perfectly clear at the outset that I fully realize the hazardous nature of any attempt to estimate the quantity of iron-ore or any other mineral in the ground. I should not have had the temerity to venture upon an inventory of the iron-ore supplies of the United States except for the requirements of the Conservation Commission. During the preparation of this inventory I have been impressed by the wide diversity in estimates of those best qualified to. hold an opinion, and I am therefore prepared in advance for criticism of my figures. Inasmuch as this report was prepared in large part for unprofessional readers, I considered it necessary to classify and describe the commercial iron-ores in a somewhat elementary manner, in order to explain the reasons for the large element of uncertainty in the estimates. I have therefore given a chemical and a geological classification of iron-ores, which place the commercial deposits in six classes: (1) magmatic segregations in basic igneous rocks; (2) contact-deposits formed in connection with igneous intrusions; (3) concentration-deposits; (4) replacement-deposits; (5) bedded-deposits, and (6) gossan-deposits. The iron-ores are further distinguished under two classes available and not available. This classification is in large measure arbitrary, and it, is evident that it will vary front time to time. In the available class are placed those ores which can be worked at a profit under conditions essentially as they exist at present. Since it depends wholly on the question of costs, the various factors which affect cost will determine the classification of any particular deposit. Actual production, * Abstract of Report prepared for the Conservation Commission. t Chief Geologist, United States Geological Survey. .

Apr 1, 1909

By Frank Milliken

IN September 1937, the Ohio Copper Co. inaugurated the treatment of its copper-bearing mill tailings at Lark, Utah. These tailings had been accumulated during the regular operation of the Ohio Copper mine and mill over the period 1007 to 1019. Gravity concentration had been employed except for a comparatively short period in 1919, when flotation was used. The ore milled during that period assayed about I per cent Cu. The chief copper mineral, chalcocite, was dissemi-nated in a shattered quartzite, mainly along fracture or cleavage planes. The quartzite gangue was extremely hard and grinding to liberation resulted in excessive sliming of the "sooty" chalcocite, with consequent high tailings losses. Recovery on the seven million tons treated was prob-ably less than 6o per cent of the total copper content. As early as 1918, attempts had been made to sample the dump. In 1928, a com-plete sampling program was undertaken. The sampling consisted of seventy-four 5-in. post auger holes drilled to the full depth of the dump at regular intervals of 250 ft. Each 10-ft. portion thereof com-prised a sample for analysis. Each sample was screened, rolled, divided into three parts, dried and assayed. The reject was retained for testing purposes. From the depth of the auger holes a contour map of the dump was plotted, Fig. 1, which gives percentages of tonnage and copper content.

Jan 1, 1940

The State School of Mines, University of Utah, is this year giving-a four weeks' course for prospectors-Jan. 8 to Feb. 3, 1917-covering the fundamentals of geology, mineralogy, mining and the metallurgical treatment of ores. There will be 36 lectures of 1 hr. each and about 20 laboratory periods, each 3 hr. in length. It is the purpose of the School of Mines to make the course entirely a practical one, and purely theoretical considerations will be omitted. The laboratory work is designed to supplement the lectures and-to give those taking the course a working knowledge of the principles involved. In order to bring the course within the reach of all, no laboratory fees will be charged, and the only fee will be a nominal one for registration in the University, of $1.00, but it is expected that those who attend will be genuinely interested in the work.

Jan 2, 1917

By Oscar Weiss

An aerial magnetometer survey was carried out by the author's geophysical organization over the Vredefort dome, where Witwatersrand beds are wrapped around a granite plug 25 to 30 miles in diameter. The well- known magnetic shales of the Lower Witwatersrand system (to be referred to as LWW) give strong negative anomalies over the out-cropping portion of the dome. This is unusual as these same shales -cause positive anomalies over large areas near Johannesburg and along a strike of 130 miles between the East Rand and Klerksdorp. It is suggested that the negative polarization is caused by heat and stresses connected with the doming. This negative polarization of sedimentary beds, which usually cause positive anomalies, is a good example of how "magnetic lows" can be caused by "structural highs." Significant magnetic anomalies should be tested by gravity and seismic methods irrespective of the sign of the anomalies.

Jan 1, 1949

By John Griffen

THEORY HYDRAULIC classification as explained by Rittinger and others was largely restricted to conditions wherein the free-falling velocities of the particles were conceived as governing the separations effected (see chapter 9). The laws governing classification under free-falling conditions have been summarized by Chapman and Mott as fo1lows:l Nomenclature r represents, in a particle of irregular shape, the round-hole screen through which the particle would pass, or some other linear dimension of the particle equivalent to d, the diameter of a sphere. s, the specific gravity of the particle. 1, the specific gravity of the water used as a medium. K, a constant. W, the velocity of the water. In an Upward Current of water 1. A particle will move upward if [V > K dr(s - 1)] and, by suit- ably regulating the speed of the water current, the larger and heavier particles may be made to fall downward against the current, leaving the smaller, lighter particles in suspension in the current. 2. Particles such that [- = -] cannot be separated in an upward- current classifier with any one velocity of current. Coal particles having a specific gravity of 1.35 cannot be separated from refuse particles of [2.5 - 1 2.5 sp. gr.] unless the ratio of diameters is less than [ 4.3 to 1, 1.35 – 1] and the raw coal must be screened to size limits within this ratio if they are to be separated. Cone-shaped vessels or vessels of other shapes that allow the use of different velocities of current at different points in the flow are used to modify this limitation.

Jan 1, 1943

By Chester A. Rowland

INTRODUCTION Comminution is generally a feed preparation step for subsequent processing stages; exceptions being when a final product such as aggregates, specification sand, Portland Cement, and similar products is produced. Grinding, the fine product phase of comminution, requires a large capital investment and frequently is the area of maximum usage of power and wear resistant materials. Grinding is most frequently done in rotating drums utilizing loose grinding media, lifted by the rotation of the drum, to break the ores in various combinations of impact, attrition and abrasion to produce the specified product. Grinding media can be the ore itself (autogenous grinding -- primary and secondary), natural or manufactured non-metallic media (pebble milling) or manufactured metallic media -- steel rods, steel or iron balls. This chapter will discuss general mill design and the specific design and application of the following types of tumbling grinding mills. [Overflow Rod Mills Figure 1 Peripheral Discharge Rod Mills Figures 2 and 3 Compartment Mills Rod and Ball Figure 6 Ball Figure 6a Pebble Mill Figure 6b Overflow Ball Mills Figure 8 Diaphragm (Grate Discharge) Ball Mills Figure 9] I. GENERAL MILL DESIGN A. Liners The interior surface of grinding mills exposed to grinding media and/or the material being ground are protected from wear and corrosion by rubber, metallic, a combination of rubber and metallic, or non-metallic wear resistant materials. B. Drives Economics at the time of plant design and mill purchase determine the drive to be used. The simplest drive is the low speed synchronous motor with speeds in the range of 150 to 250 RPM connected to the mill pinion-shaft by either an air clutch or flexible coupling.

Jan 1, 1982

By Arthur S. Huey

AS the end of the collegiate year approaches, the problem of student employment again becomes acute. This subject divides itself into two phases: (1) employment prior to graduation and (2) employment following graduation. "Experience is the best teacher," according to the old adage, and the student needs all that he can get. He is afforded every opportunity to reap from his books the theories and principles of scientific and engineering practice, and occasionally he is privileged to make field trips to observe the practical side of the subject. But it is not enough for one to know the theory and to observe its application; it is also desirable for the student to get the feel of the tools during his training Such opportunities can only be afforded the student if the major companies cooperate with the universities in giving the student part-time and vacation work.

Jan 1, 1932

By Gustav S. Preller, A. M. Gaudin

For the past 20 years it has been generally accepted that the flotation process is made possible as a result of the action of certain chemical substances on the surface of the mineral particles. In fact, it has been suspected, but never demonstrated, that this action produces on the mineral particles a layer of chemosorbed ions that is of the order of magnitude of a monolayer; that is, of a layer one ion deep. It is the purpose of this paper to present certain initial results on the relationship of mineral surface area and collecting-agent requirements. Up to about eight years ago, all considerations involving the surface area of finely divided solids were limited by the fact that there was not in existence any general method for determining the extent of such surfaces, especially on mixed materials. However, the method invented by Brunauer, Emmett and Teller1 has changed the situation by providing a general procedure (BET method). This procedure has already been applied to mineral dressing problems2-4 and has now received confirmation as to accuracy by an entirely independent method.' This confirmatory or "absolute" method uses a very tedious calorimetric procedure and is not suitable for routine determinations; but still another method6 now is available to relate the results obtained by the BET method to those obtained calorimetrically. This Harkins-Jura method has the advantage that it does not require an assumption concerning the parking area per molecule of adsorbate (nitrogen). In the original work by Brunauer, Emmett and Teller the parking area per nitrogen molecule was calculated to be either 13.8?2 or 16.2?2. according as the nitrogen was assumed to be closely packed as in crystalline nitrogen or loosely packed as in liquid nitrogen. The initial experimental work from this laboratory2 supported the 13.8?2. figure; that work was done on glass spheres of measurable geometric area and it was based on the assumption that the glass spheres were truly smooth; any deviation from this assumption would, of course, have made the calculated parking area higher. Harkins and Jura's results suggest that the parking area is not the same for all substrata; furthermore, the average of the 87 measurements cited indicates a "preferred" parking area for the nitrogen molecule of about 15.2?2., with extreme values at 13.7 and 16.9?2. The average of 21 determinations made in connection with this study7 gave a parking area of 14.3 ± 0.27?². In view of this result and of the Harkins and Jura

Jan 1, 1947

By Gustav S. Preller, A. M. Gaudin

For the past 20 years it has been generally accepted that the flotation process is made possible as a result of the action of certain chemical substances on the surface of the mineral particles. In fact, it has been suspected, but never demonstrated, that this action produces on the mineral particles a layer of chemosorbed ions that is of the order of magnitude of a monolayer; that is, of a layer one ion deep. It is the purpose of this paper to present certain initial results on the relationship of mineral surface area and collecting-agent requirements. Up to about eight years ago, all considerations involving the surface area of finely divided solids were limited by the fact that there was not in existence any general method for determining the extent of such surfaces, especially on mixed materials. However, the method invented by Brunauer, Emmett and Teller1 has changed the situation by providing a general procedure (BET method). This procedure has already been applied to mineral dressing problems2-4 and has now received confirmation as to accuracy by an entirely independent method.' This confirmatory or "absolute" method uses a very tedious calorimetric procedure and is not suitable for routine determinations; but still another method6 now is available to relate the results obtained by the BET method to those obtained calorimetrically. This Harkins-Jura method has the advantage that it does not require an assumption concerning the parking area per molecule of adsorbate (nitrogen). In the original work by Brunauer, Emmett and Teller the parking area per nitrogen molecule was calculated to be either 13.8?2 or 16.2?2. according as the nitrogen was assumed to be closely packed as in crystalline nitrogen or loosely packed as in liquid nitrogen. The initial experimental work from this laboratory2 supported the 13.8?2. figure; that work was done on glass spheres of measurable geometric area and it was based on the assumption that the glass spheres were truly smooth; any deviation from this assumption would, of course, have made the calculated parking area higher. Harkins and Jura's results suggest that the parking area is not the same for all substrata; furthermore, the average of the 87 measurements cited indicates a "preferred" parking area for the nitrogen molecule of about 15.2?2., with extreme values at 13.7 and 16.9?2. The average of 21 determinations made in connection with this study7 gave a parking area of 14.3 ± 0.27?². In view of this result and of the Harkins and Jura

Jan 1, 1947

By E. T. Turkdogan, L. J. Martonik, R. J. Fruehan

Electrochemical measuretnents of the solid oxide electrolyte galvanic cells CY-Cr2O3 I ZrO2 (CaO) 1 O (in Fe alloy) CY-Cr2O3 I Tho2 (Y2O3)I O en Fe alloy) have been made at 1600°C (2912°F) in order to test the Performance of such cells at liquid steel temperatures. The oxygen pvobe (cell) consists of a disk of ZrO2 (CaO) or Tho2 (Y2O3) electrolyte fused at one end of a silica tube filled with a mixture of Cr-Cr2O3 which is the reference electrode. Upon immersion in liquid steel, the electromotive force readings achieve a steady value within a few seconds, and remain steady for 30 win or more. The perforwzance of the probes has been tested using Fe-O, Fe-Si-O, Fe-Cr-O, Fe-V-O, and Fe-Al-O alloys; the oxygen contents of liquid steel derived from the measured electromotive forces are in satisfactory agreement with those determined by arulysis. Use of the probe in the deoxi-datiorz of steel, in laboratory experiments, is discussed. The results indicate that there is insignificant electronic conductivity in ZrO2(CuO) at oxygen activities down to those corresponding to 10 ppm in steel. At lower oxygen activities, probes tipped with ThOn (Y2O3) disks perform satisfactorily at oxygen activities down to 1 ppm O or less. THE key to the control of deoxidation of steel is a sensing device to measure rapidly the concentration of oxygen in liquid steel in the furnace, ladle or tun-dish at any desired stage of deoxidation. The analysis of the cast steel by the neutron-activation or vacuum-fusion method gives total oxygen as oxide and silicate inclusions. This analysis is important for guidance to steel cleanliness; however, such a postmortem is of little value in the control of deoxidation of liquid steel. At the General Meeting of the American Iron and Steel Institute in New York, 1968, Turkdogan and Fruehan' presented a paper on the preliminary results of the work done in this laboratory on rapid determination of oxygen in steel by an oxygen probe. Details of the work done in this laboratory leading to the development of a galvanic cell for the determination of oxygen in liquid steel, and the results of the tests made are given in this paper. It was through Wagner's contributions, since the early Thirties, to the physical chemistry of semiconductors in general that it ultimately became possible to construct galvanic cells for application at high temperatures. In 1957, Kiukkola and wagner2 successfully demonstrated the use of several solid electrolytes in measuring the free energies of several chemical reactions, in particular, the use of lime-stabilized zir-conia in high-temperature oxidation reactions. Starting 7 years later, a number of papers appeared in the technical literature3-' demonstrating possible applicability of galvanic cells for the determination of oxygen in liquid steel. In the earliest work, Japanese investigators3j4 experimented with various types of reference electrodes, e.g., graphite-saturated liquid iron at 1 atm CO or Ni-NiO mixtures; the results obtained, though promising, were not of sufficient accuracy. Except for the work of Baker and West,6 all other investigators5,7,8 showed that ZrO2(CaO) electrolyte could be used for this purpose. The main part of the galvanic cell used by Fischer and ~ckermann' and by schwerdtfeger7 (the latter work was done in this laboratory), consisted of a ZrO2(CaO) tube, -1 cm ID, closed at one end, with a platinum contact wire fixed mechanically inside the closed end. The tube was flushed with a gas of known oxygen partial pressure, e.g., air, CO-CO2 or H2-CO2 mixtures; gas along with the platinum lead wire served as the reference electrode. The oxygen contents derived from measured electromotive forces agreed reasonably well with the oxygen contents determined by vacuum-fusion analysis. It is evident from recent investigations that the electromotive force technique using a solid oxide electrolyte is fundamentally well suited for the determination of oxygen in liquid steel. However, it is equally clear that the cell arrangement of the type as commonly used is in need of considerable improvement, as it exhibits several shortcomings for industrial and even laboratory use. 1) Because of its size, the zirconia tube, though stabilized, has a poor resistance to thermal shock. 2) Fine pores and microcracks, which are invariably present in zirconia tubes, are detrimental to the satisfactory operation of the cell, particularly when gas reference electrodes are used. 3) Air or carbon dioxide reference electrodes give rise to high electromotive force readings; as a result, the determination of oxygen in steel becomes less accurate. For higher accuracy, the oxygen partial pressure of the reference electrode should be in the range similar to that of oxygen in steel. 4) Even in laboratory experiments, difficulties are experienced when flushing the tube with gases and maintaining the proper gas flow rate. Fischer and Ackermann,' who used air as the reference electrode, reported that when the flow rate was too low, furnace gases would leak into the electrolyte tube, therefore lowering the oxygen potential and measured electromotive force. The required flow rate in order to avoid leakage depended on the tightness of the electrolyte tube which varied with different tubes, thus making it difficult to predict in advance the required flow rate. However, if the flow rate is too high the inside wall of the electrolyte tube would be cooler than the wall

Jan 1, 1970

By Gustav S. Preller, A. M. Gaudin

Fox the past 20 years it has been generally accepted that the flotation process is made possible as a result of the action of certain chemical substances on the surface of the mineral particles. In fact, it has been suspected, but never demonstrated, that this action produces on the mineral particles a layer of chemosorbed ions that is of the order of magnitude of a monolayer; that is, of a layer one ion deep. It is the purpose of this paper to present certain initial results on the relationship of mineral surface area and collecting-agent requirements. Up to about eight years ago, all considerations involving the surface area of finely divided solids were-limited by the fact that there was not in existence any general method for determining the extent of such surfaces, especially on mixed materials. However, the method invented by Brunauer, Emmett and Teller1 has changed the situation by providing a general procedure (BET method). This procedure has already been applied to mineral dressing problems2 -4 and has now received confirmation as to accuracy by an entirely independent method.5 This confirmatory or "absolute" method uses a very tedious calorimetric procedure and is not suitable for routine determinations; but still another method6 now is available to relate the results obtained by the BET method to those obtained calorimetrically. This Harkins-Jura method has the advantage that it does not require an assumption concerning the parking area per molecule of adsorbate (nitrogen). In the original work by Brunauer, Emmett and Teller the parking area per nitrogen molecule was calculated to be either 13.8Å2. or 16.2Å2. according as the nitrogen was assumed to be closely packed as in crystalline nitrogen or loosely packed as in liquid nitrogen. The initial experimental work from this laboratory2 supported the 13.8Å2. figure; that work was done on glass spheres of measurable geometric area and it was based on the assumption that the glass spheres were truly smooth; any deviation from this assumption would, of course, have made the calculated parking area higher. Harkins and Jura's results suggest that the parking area is not the same for all substrata; furthermore, the average of the 87 measurements cited indicates a "preferred" parking area for the nitrogen molecule of about 15.2Å2., with extreme values at 13.7 and 16.9Å2. The average of 21 determinations made in connection with this study' gave a parking area of 14.3 ± 0.27Å2. In view of this result and of the Harkins and Jura

Jan 1, 1946

By L. E. Shiffman

W. J. Parton—Those operators faced with the problem of treating fine coal whether in bituminous or anthracite will find this paper most timely. I would like to take this opportunity of discussing Mr. Schiffman's paper and at the same time express certain views relating to our Tamaqua plant. I would like to ask the author what type of impeller and diffuser is used in the Denver cells? Screen analysis of products from individual cells indicate that coarser material resists flotation and only floats after greater retention time in the last few cells. Also, the need for a scavenger screen to reclaim non-floated coal particles further stresses this point. I have always felt that more efficient means of cleaning coal between 10-mesh and 28-mesh existed than flotation. Reagent and power costs are high for the flotation process. When floating +28-mesh particles, cell capacity is lowered and some of the particles are lost with the refuse. The Tamaqua plant of the Lehigh Navigation Coal Co. floats —28-mesh coal and capacity of recoverable coal is 40 tph for 1800 cu ft of Denver cells; or 0.05 tons per cu ft of cell. At Kimberly 7.75 tph for 600 cu ft of cell gives 0.013 ton per cu ft of cell. At Bessie 14 tph for 800 cu ft of cell gives 0.017 tons per cu ft of cell. It would be appreciated if the author would comment on what he feels is the upper size limit of particle to attain most efficient utilization of the flotation process. The dewatering screw is a very interesting development since it offers a simple way to prepare coal sludge for more complete clewatering by drainage or mechanical dewatering on screen or filters. In other words it could be used to accomplish the same thing as a thickener tank. I would appreciate having the author's comment on how he thinks such a screw dewaterer would work on a froth.* The process as used in floating coal at the Bessie and Kimberly plants may be referred to as more of a bulk oil float in contrast to a froth flotation process. Experiments on increasing capacity of cells is most interesting since we are going through such an experi- mental period at the present time. Recently a double overflow was installed on our No. 30 Denver cells. So far results are not conclusive. In reviewing this paper the following comments are made pertaining to investigation of methods for increasing capacity: Supercharging: Supercharged air in matte flotation or for that matter the use of the normal amount of air drawn in by the impeller would in all probability cause such an aeration in the cell as to destroy the buoyant effect given to the coal particles by the excessive amount of kerosine used. In other words, air creates an agitation zone throughout the cell, creating a boiling and thereby giving a lower recovery in the cell. It would be interesting to know whether the 7 pct increase in recovery was with no air being admitted to the stand pipe. Changing Impeller Speed: The speed of a receded disc impeller for a No. 30 cell as recommended by the Denver Equipment Co. is, I believe, approximately 250 rpm. At this speed and using supercharged air in excess of 8-oz pressure, we have observed a boiling action in the cells. In our flotation we endeavor to obtain some degree of froth flotation using pine oil as a frother. The boiling action as caused by increasing the amount of air added to the cells is detrimental to recovery in froth flotation. It is our belief that to obtain increased recovery from a cell in froth flotation, additional air must be introduced but at the same time this air must be dispersed throughout the pulp in the form of small bubbles and this can only be done by increasing the speed of the impeller. Therefore, if Mr. Schiffman decreased the speed of the No. 30 impellers and at the same time continued to use supercharged air, the boiling action may have been increased because larger bubbles developed. The lower recovery as reported could be due to this factor. Decreasing the impeller speed will definitely decrease the power consumed but may have other disadvantages. First, we believe it will permit "sanding" in the cell and this in our opinion will increase the wear on the impeller and diffuser, especially so, if there is pyrite and/or sand present in the feed. "Sanding" in the cell when air is used, as in froth flotation, will effect the dispersion of this air and cause boiling.

Jan 1, 1951

By L. E. Shiffman

W. J. Parton—Those operators faced with the problem of treating fine coal whether in bituminous or anthracite will find this paper most timely. I would like to take this opportunity of discussing Mr. Schiffman's paper and at the same time express certain views relating to our Tamaqua plant. I would like to ask the author what type of impeller and diffuser is used in the Denver cells? Screen analysis of products from individual cells indicate that coarser material resists flotation and only floats after greater retention time in the last few cells. Also, the need for a scavenger screen to reclaim non-floated coal particles further stresses this point. I have always felt that more efficient means of cleaning coal between 10-mesh and 28-mesh existed than flotation. Reagent and power costs are high for the flotation process. When floating +28-mesh particles, cell capacity is lowered and some of the particles are lost with the refuse. The Tamaqua plant of the Lehigh Navigation Coal Co. floats —28-mesh coal and capacity of recoverable coal is 40 tph for 1800 cu ft of Denver cells; or 0.05 tons per cu ft of cell. At Kimberly 7.75 tph for 600 cu ft of cell gives 0.013 ton per cu ft of cell. At Bessie 14 tph for 800 cu ft of cell gives 0.017 tons per cu ft of cell. It would be appreciated if the author would comment on what he feels is the upper size limit of particle to attain most efficient utilization of the flotation process. The dewatering screw is a very interesting development since it offers a simple way to prepare coal sludge for more complete clewatering by drainage or mechanical dewatering on screen or filters. In other words it could be used to accomplish the same thing as a thickener tank. I would appreciate having the author's comment on how he thinks such a screw dewaterer would work on a froth.* The process as used in floating coal at the Bessie and Kimberly plants may be referred to as more of a bulk oil float in contrast to a froth flotation process. Experiments on increasing capacity of cells is most interesting since we are going through such an experi- mental period at the present time. Recently a double overflow was installed on our No. 30 Denver cells. So far results are not conclusive. In reviewing this paper the following comments are made pertaining to investigation of methods for increasing capacity: Supercharging: Supercharged air in matte flotation or for that matter the use of the normal amount of air drawn in by the impeller would in all probability cause such an aeration in the cell as to destroy the buoyant effect given to the coal particles by the excessive amount of kerosine used. In other words, air creates an agitation zone throughout the cell, creating a boiling and thereby giving a lower recovery in the cell. It would be interesting to know whether the 7 pct increase in recovery was with no air being admitted to the stand pipe. Changing Impeller Speed: The speed of a receded disc impeller for a No. 30 cell as recommended by the Denver Equipment Co. is, I believe, approximately 250 rpm. At this speed and using supercharged air in excess of 8-oz pressure, we have observed a boiling action in the cells. In our flotation we endeavor to obtain some degree of froth flotation using pine oil as a frother. The boiling action as caused by increasing the amount of air added to the cells is detrimental to recovery in froth flotation. It is our belief that to obtain increased recovery from a cell in froth flotation, additional air must be introduced but at the same time this air must be dispersed throughout the pulp in the form of small bubbles and this can only be done by increasing the speed of the impeller. Therefore, if Mr. Schiffman decreased the speed of the No. 30 impellers and at the same time continued to use supercharged air, the boiling action may have been increased because larger bubbles developed. The lower recovery as reported could be due to this factor. Decreasing the impeller speed will definitely decrease the power consumed but may have other disadvantages. First, we believe it will permit "sanding" in the cell and this in our opinion will increase the wear on the impeller and diffuser, especially so, if there is pyrite and/or sand present in the feed. "Sanding" in the cell when air is used, as in froth flotation, will effect the dispersion of this air and cause boiling.

Jan 1, 1951

By E. R. Borcherdt

IT is notable that the first large-scale mine operation equipped entirely with detachable bits was the Badger State mine of the Anaconda Copper Mining Co. in Butte, Montana, just 30 years ago. This mine in 1922 was producing approximately 1200 tons of ore per day. Much of the data presented in C. L. Berrien's article' describing the development and installation of the Hawkesworth detachable drill bit were obtained from these operations. As in any pioneering effort, no precedent existed and many difficult problems required solution, so that the changeover to detachable bits at all Butte hill mines was not completed for 6 years. There was widespread disbelief as to the probable efficiency of the new installation. Some attempts were made in 1931 by the owners of the Hawkesworth patents to interest Ontario gold mine operators in the bit. These efforts were not successful, but they undoubtedly stimulated thinking which resulted in the invention and patenting of several well-known Canadian detachable bits, one of which is now a widely used throwaway bit. The success of the Butte installation also led to the development of the threaded type of bit connections by several well-known manufacturers, and in 1935 these bits were introduced to the mining industry on a national scale. The original Hawkesworth bit was not provided with a water hole but, depended upon water passing through the clearance opening between the tongue in the bit and the groove in the rod to flush cuttings from the drill hole, see Fig. 1. In December 1935 it was found that this method of introducing drilling water to the bit face resulted in high dust counts. To correct this a water hole was drilled on the central axis of the bit, passing through the tongue. Unfortunately, quenching water would rise through the small water hole, spot-hardening the tongue to cause breakage, never completely eliminated. In the fall of 1936 large-scale tests indicated that savings would be effected by use of a threaded type of bit, which was therefore adopted as standard for all Butte mines. This type of bit was used until 1947, when it was superseded by a one-use slip-on type. Since the first use of the Hawkesworth bit every detachable bit of importance has been investigated, and where advantages which might reduce costs or increase efficiency were indicated, substantial tests of the bit were carried on in the Butte mines. When tests demonstrated the advisability of changing from one kind of detachable bit to another the change was made at one level or in one area each day until the new rod and bit equipment was used throughout the mine. This involved a minimum of cost and disruption of drilling. Intelligent selection of a detachable bit to obtain optimum results requires careful consideration to achieve a balance between the three principal types of equipment used in the drilling process: 1—drill bits, 2—drill steel, and 3—drilling machines. Optimum results imply maximum output and minimum cost per unit of output. Since every rock type differs in drillability and it is generally impractical to provide equipment for more than one or two types of rock which may occur in one operation, selection of equipment must encompass average drilling conditions. However, on exceptional occasions several widely differing conditions may make it mandatory to provide equipment best suited to each condition. The choice of rock-drilling equipment is a most controversial subject and one that is further complicated by unreliable and frequently misleading performance claims. Small operators without the means for making accurate evaluations of equipment frequently suffer from these over-enthusiastic claims. It is apparent from experience in rock drilling throughout the world that rock drillability is not alike in any two places, and that selection of proper equipment can only be made after conducting thorough trials of various types of equipment. Some recent drilling tests in tactite and hornstone at the Darwin, California mine of the Anaconda Co. present some interesting clues on rock drillability. Microscopic examination of thin sections of these rocks reveals that mineral composition and rock texture are equally important in governing drillability. The Darwin hornstone is at times so abrasive that the carbide bit cutting edges become flattened to 3/32 in. in 2 to 4 ft of drilling, and some carbide bits were dulled to this point after 9 to 10 in. of drilling. This wear was determined to be the proper point for resharpening to eliminate carbide insert breakage or breakage of the steel rod when drilling with 1½ to 1?-in. bits, with a drifter of 2 3/4-in. diam and 90 to 100 psi air pressure, see Supplement A. Before considering the merits of various bit designs it may be well to review the mechanics of drilling rock with percussion drills. A sharp bit cuts by penetration and chipping. The amount of penetration governs the amount of chipping and depends upon the contact area of the cutting edge, the foot-

Jan 1, 1954