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By James K. Stanley

When a plastically deformed mctal is heated to a certain temperature, it undergoes a complete change in microstructure, the consequence of which is a marked alteration of mechanical properties such as hardness, strength and ductility. This change in microstructure occurs by re-crystallization and is a process involving the formation and growth of strain-free nuclei in the deformed matrix. These nuclei continue to form with time and grow until a new grain structure replaces the cold-worked structure. The formation of the nuclei in the deformed sheet of I per cent silicon ferrite is illustrated in Fig. I. If a recrystallization is conducted iso-thermally, a nleasurement of the fraction of the matrix recrystallized plotted against time leads to an isothermal recrystallization curve, first established in 1942.1 Such recrystallization curves are shown in Fig. 4. These curves show that the recrystallization of a cold-worked metal is best considered as a rate process rather than a time independent process such as is suggcsted by the improperly named and misleading recrystallization diagrams involving deformation, temperature, and grain size.2-5 Any study of the rate of recrystallization resolves itself into a study of the component rates of nucleation and growth. The previous paper1 provided an account of the determination of the rates of nucleation; N, the number of nuclei formed per unit time per unit of unre-crystallized area, and growth, G, the radial growth in unit time, in the recrystallization of silicon ferrite (I per cent Si). A method was developed to furnish experimental data for evaluating the effect of the recrystallization variables on N and G. The mathematical analysis for the derivation of the isothermal recrystallization curve was developed for a two-dimensional case (in thin sheets where the diameter of the recrystallized grain is greater than the sheet thickness). The present paper is a continuation of the study of recrystallization of I per cent silicon ferrite begun in 1942.1 In this work the effect of the factors deformation, temperature, grain size, composition, and recovery is studied with respect to the rates of liucleation and grouth. At the same time it is shown how the cooperation of N and G affects the isothermal recrystallization curve. The study of these variables on N and G has a definite bearing on the mechanism of recrystallization and should lay a foundation for a rational theory on the subject. Procedure The procedure used in these experiments was essentially the same as that described previously but in certain details some changes were made in the interests of speed. With few exceptions, the procedure is the same for the study of all the variables. Where the procedure is different from standard manipulations attention is called to this fact.

Jan 1, 1945

By Michael Tenenbaum

Selected characteristics of nonmetallic phases were determined in low-carbon sheet steels which had been produced by various refining; deoxidizing, and casting procedures. The results showed little influence of the refining process on the quantity, distribution, size, and character of the nonmetallic phases. The quantity of inclusions was governed by the deoxidation procedures. While the mode of casting influenced the internal structure and the pattern of inclusion distribution, it had little effect on the quantity of nonmetallics. The vast majority of inclusions were under 30 µ in diam, and there was little change in size distribution during hot and cold rolling. The composition and structure of the inclusions were expained an the basis of known equilibrium phase diagrams for the appropriate systems. Additional work is needed to explain some of the observations of this paper. HENRY Marion Howe's contributions to the metallurgical profession include his many writings, the wisdom that he imparted to his students, and the inspiration that he gave to the many scholars, teachers, and industrialists who knew him personally. It was Professor Howe's intimate colleague, Albert Sauveur,1 who, in the first Howe Memorial Lecture, could have been speaking for the metallurgist of our generation as well as his own when he wrote, "...how many of us on the point of going astray were called back to the right path by Howe's guiding torch?" More than a quarter of a century later, Francis Foley,2 one of the last in the Howe Lecture series who worked directly with the great man, delivered the ultimate tribute, "...truth, in its time is so hard to identify." While Professor Howe strove to be right, "he was content if he provoked the reader to take the next step along the path to greater knowledge." Like so many of my predecessors, I sought guidance for this lecture by referring to Professor Howe's two classic texts, Metallurgy of steel,3 and Metallography of Steel and Cast Iron: the former book interpreting the facts as he saw them near the start of his career, and the latter—published 26 years later—pointing to many fertile fields for metallurgical investigation. Those of us who harvest the fruits of Henry Marion Howe's contributions to the metallurgical profession now look with wonder at the tremendous legacy he has provided. It is rewarding to study his writings and to recognize the anticipation of such aspects of modern ferrous metallurgy as: 1) Direct reduction 2) Steelmaking reactions 3) The approaches and benefits of degassing 4) Solidification mechanisms 5) Habits and behavior of nonmetallic inclusions 6) Recrystallization phenomena 7) Slip and slip propagation 8) Dislocations 9) Discontinuous deformation 10) Microscopic and macroscopic segregation

Jan 1, 1970

By J. H. Tu

The technical paper correctly points out that the kriging variance is not a good measure of the uncertainty of the estimated (i.e., kriged) value of individual blocks. The authors claim that their proposed jack-knife method, which is a rekriging of each block by eliminating, in turn, one sample from m the original sample set and then taking the average of the rekriged estimates, not only gives good block estimates, but the resulting jackknife kriging standard deviation is a useful indicator of the "true uncertainty associated with block estimates." However, they immediately abandon the idea of using the block-by-block standard deviations, reasoning that these standard deviations are not independent and that there is no easy way to utilize them. There may be another reason for not using them. The jackknife standard deviations for individual blocks given in their example are mostly in the range of 0.004 to 0.005 oz/st (0.14 to 0.17 g/t) with only one block having a high value of 0.012 oz/st (0.41 g/ t). These individual block standard deviations are as low as the jackknife standard deviation for the mean grade of the entire shape, i.e., 0.0041 oz/st (0.14 g/t). Do they represent the "true uncertainty" .of the individual block estimates? Could the authors explain this? In a global shape consisting of a large number of blocks, any given sample will affect the kriged estimate of only those few blocks within its vicinity. This is the rationale for the authors' selective rekriging, making the jackknife algorithm more efficient. On second thought, why not do away with jackknifing altogether? Just cumulate and normalize, if necessary, the kriging weights of each sample used during the ordinary block kriging process, and then compute the global variance from these kriging weights and their respective sample grades? After all, isn't the global mean grade nothing but the weighted average of the samples used in the estimation? Reply by G.S. Adisoma and M.G. Hester The jackknife is one of the many tools in a practitioner's toolbox to solve estimation problems. The strengths of the technique lies in its simplicity, i.e., it uses the concept of mean and standard deviation and the fact that it can be easily combined with other tools, in this case kriging. Because the jackknife kriging (JK) estimate is also the mean of the pseudovalues, the JK standard deviation is attractive just as the standard deviation of the mean explains the variability of the data. The difference is that the pseudovalue calculation in jackknife kriging uses the ordinary kriging (OK) weighting scheme instead of simple arithmetic averaging. The data used to illustrate the jackknife technique in the paper con¬sist of high values that are roughly three times the low values. The resulting JK estimate of the block grades show that the highest estimate is roughly twice the grade of the lowest estimate. The contrast between the low and the high estimate is more evident in the JK estimate than in the OK estimate, even though the mean grades of the blocks for the two estimates are very similar. Nonetheless, in this paper, we are concentrating more on the need for a more realistic measure of uncertainty, or precision, for the estimate. Unlike its OK counterpart, the JK standard deviation of the blocks clearly reflects the original data variation. The highest JK standard deviation of the blocks is three times its lowest value. This follows our intuition that, when the samples used to estimate a block is more variable, the resulting estimation variance (or standard deviation) should be higher than the case where the samples are more uniformly valued. However, block-by-block standard deviation or variance is of little practical value in reserve estimation and classification, as well as in mine planning. One is usually more interested in quantifying not the variance of the individual block estimate, but the uncertainties associated with a much larger dimension, such as the minable reserve. Thus, the thrust of the paper is to find a simple way to obtain a single estimation variance or standard deviation associated with the reserve grade estimate. The discussion by J.H. Tu did not mention how one would obtain the global variance from the OK weights and the sample grades. As a technique that offers a data value-based measure of uncertainty for its estimate, the "leave-one-out" jackknife fills this need nicely through the block kriging shortcut approach described in the paper. Note: The first column and the last two columns of Table 3 in the paper should have contained a single number each, namely, an OK estimate of 0.0317, a JK estimate of 0.0333 and a JK standard deviation of 0.0041 oz/st, respectively, for the shape, as are obvious from the text.

Jan 1, 1997

By Hans Ehdwall

INTRODUCTION Investigations of radon and radon daughter concentrations in Swedish [non-uranium] mines started in the late 1960's. The first screening measurements showed that the average annual exposure to radon and radon daughter products was 4.7 WLM. The main reason for high radon and radon daughter concentrations was inefficient ventilation and radonrich water entering the mine. In the radon regulations worked out later it was stated that no miner should be exposed to more than 60 000 pCi h/1 equilibrium equivalent concentration of radon annual exposure, corresponding to 3.6 WLM. Now, 1981 the situation has changed considerably. From the average annual exposure of 4.7 WLM in 1970 it is now only 0.7 WLM. Sweden has up to now had only one [uranium] mine and the work there has only been investigative. However, there are plans for a commercial uranium mine in another part of Sweden. The radon problems in these mines are widely different depending on the mineralogy. NON-URANIUM MINES The radiation problems in Swedish mines were not recognised until the late 60's. The first radon and radon daughter measurements were made in some sulphide ore mines in 1967 (1). The radon and radon daughter concentrations were surprisingly high for non-uranium mines. In order to have a complete picture of the radon situation in Swedish mines the National Institute of Radiation Protection (NIRP) decided to make measurements in all, at that time about 60 mines (2). To get results as fast as possible measurements on radon gas seemed most appropriate to start with. Sampling was made by mailing a number of evacuated 4.8 litre conventional propane containers from NIRP to each mine. The containers were then opened at the place of interest. After sampling the containers were sealed and then mailed back to the institute for measurement. The measurements were made in ionization chambers. This method only gave the radon concentration and the radon daughter concentration was estimated by multiplying the radon concentration by an assumed equilibrium factor. The equilibrium factor is defined as the ratio of the total potential alpha energy for the given daughter concentration to the total potential alpha energy of the daughters if they are in equilibrium with the given radon concentration. The results of this first preliminary survey indicated that a great many of the Swedish miners probably had an annual radon daughter exposure of more than 3.6 WLM. As the radiation exposure in non-uranium mines was not regulated in either the Swedish Radiation Protection Act or the Swedish Labour Protection Act work was started on special radon regulations. A lung cancer mortality study was also started. To check the results of the first survey and to get experience and knowledge of radon problems in mines, it was decided that personnel from the NIRP should visit each mine for a detailed investigation of radon and radon daughter concentrations starting with the ones with the highest radon concentrations. The main reasons for these so-called "basic measurements" were: 1. To estimate the doses received by Swedish miners 2. To find the sources of the high radon and radon daughter concentrations 3. To find appropriate counter-measures 4. To determine the most typical equilibrium factor for each mine. Unlike most uranium mines the reason for high radon concentrations in non-uranium mines is seldom the occurrence of highly radioactive minerals. The main sources were found to be waste-rock and radon-rich water. In order to filter and warm up the inlet air, especially in winter time, it was very common at that time to suck the air through broken wasterock. By doing so the air was contaminated with radon from the waste-rock and radon-rich water in it. It is noteworthy that the radium and uranium concentration in the waste-rock is relatively low. The uranium concentration is only of the order of 15 - 20 ppm. The action to prevent this contamination of the inlet air was to change the direction of the ventilation and in the case of radon-rich water entering the mine the action was to prevent the air coming into contact with the water. The first calculation of the radon daughter exposure of Swedish miners was based on radon gas measurements. The radon daughter concentration was estimated by using an assumed equilibrium factor of 0.5. Later when the mines were visited by institute staff it was possible to compare the assumed equilibrium factor with the measured ones. It was found that the factor varied from 0.15 at the air inlet to 1.0 at the air outlet and the average equilibrium factor on workplaces for almost all mines was between 0.4 and 0.6. The result of the exposure calculation in 1970 showed that more than 40 % of the miners had an annual radon daughter exposure of more than 3.6 WLM. The overall average was 4.7 WLM and the maximum annual expo-

Jan 1, 1981

By Jon C. Volkwein

The U. S. Bureau of Mines investigated a method of using inert gas to prevent the formation of explosive gas mixtures in auger highwall mining of coal. A combination of gasoline and diesel engine exhaust gases was introduced into the auger drill hole using a short section of pipe located at the collar. Gas samples were taken and analyzed on site with infrared detectors for oxygen, carbon dioxide, methane, and carbon monoxide. Evacuated bottle samples were also taken and analyzed by gas chromatography at the Pittsburgh Research Center. These gas results were analyzed for explosibility. Personal exposure to carbon monoxide was also monitored. The highest methane level observed was 9.55 pct. The Inert gas levels, (carbon dioxide and nitrogen) were sufficiently high to prevent any ignition of the methane. Results showed that for all conditions during mining, gas concentrations were non-explosive. The maximum personal time weighted average sample for carbon monoxide was 20 ppm. This system provides a safe, inexpensive, simple method for preventing explosions during auger mining. INTRODUCTION The auger highwall mining method is an effective method to recover coal from a reserve when removal of the overburden by surface mining equipment becomes uneconomical. In this method of mining, a horizontal auger enters the coal seam from the surface mine bench under the highwall and the coal is drilled in a series of parallel holes. Historically, coal mined from the surface is relatively shallow, and over time, methane associated with the coal has dissipated through the surface. In most circumstances, little methane has been found associated with auger mining. However, mining technology has enabled surface mining of deeper reserves of coal. Furthermore, environmental constraints have forced the highwall extraction method to be used to remove coal under wetlands, further increasing the chances of encountering methane. Recently incidents of methane explosions at a few auger mining operations have resulted in injuries and increased testing for methane at the collars of auger holes. The fuel source of the reported explosions was not necessarily limited to methane, but may also have involved coal dust. The Mine Safety and Health Administration (MSHA) met with the Bureau to discuss what technology might be available to enable the safe resumption of mining. The discussion included the difficulty of ventilating through the solid shafts of the augers, that steel bits probably created the ignition source, and that perhaps inerting the holes with low oxygen and high carbon dioxide concentrations from the machine's diesel exhaust was a potential solution. Considering the ventilation aspects of the problem, it was not clear If ventilation could be reliably established. If some degree of ventilation to the front of the mining head is achieved, it may combine with methane to bring the hole atmosphere from a rich, nonexplosive mixture to an explosive mixture. Furthermore, it may not prevent a dust explosion in such a mining configuration. Lack of access through the shafts of auger type mining machines further limits the ability to add water or air to cool bits to prevent an ignition source from developing. Either of these approaches would also be expensive. The process of mining coal In an inert atmosphere has been considered in the past, but to our knowledge, never implemented (Department of Interior, 1970). Clearly, implementation in underground mining would be more complicated. On a mine bench open to the atmosphere, however, adding inert gas to the mining head could provide a quick, feasible method to prevent explosions at auger highwall mining operations. Also the problem of how to move the inert gas to the cutting head of the machine had to be considered. Preventing explosions on auger mining machines using inert gas requires three primary considerations: first is the source of inert gas; second, placing the inert gas at the cutter head; and third, monitoring the hole atmosphere. Any gas source having an effective inert gas concentration of 34 volume pct or greater will prevent methane from Igniting (Zabetakis, 1965). Sources of inert gas considered for this application included liquid nitrogen, modified shipboard inert gas generators (for hydrocarbon shipping and transfer), jet turbine engine (Paczkowski, et. al., 1982), the auger's diesel engine and a gasoline engine. Operation cost, purchase cost and availability limited our testing to the diesel and gasoline engines. This work tested each engine, separately and combined. To ensure effectiveness, both company and enforcement personnel need to know how to monitor the condition of the inerted hole. Measurements at depth inside the hole are possible by remote sampling through rigid tubing, but this method is Impractical for routine monitoring. Continuous monitoring of the exhaust gas stream is an alternative. The U. S. Bureau of Mines evaluated an inert gas system at an auger mining operation at a surface mine near Owensboro, KY. Coal was mined from the Number 9 Coalbed in Henderson Co. KY. Tests were conducted in January and March of 1992.

Jan 1, 1993

By Thomas R. Smith

This paper covers oil shale resources in those countries that border the Pacific Rim. The major known resources around the Pacific Rim occur in the Western United States, Australia, the People's Republic of China, (PRC) and the Thailand/Burma region. The location of these deposits is shown in Figure 1. In 1965, the U.S. Geological Survey estimated world oil shale deposits of over 4 quadrillion tons having a potential oil yield of over 2 quadrillion barrels. If all this were extracted, it could meet the world's entire energy needs far into the future. However, the Survey also estimated the spent shale waste could cover all of the surface of the world to a depth of about 10 feet. Thus, for this and many other technical and economic reasons, it does not appear to be feasible to develop a large portion of the world's oil shale resources in this century; nor will shale in itself solve our energy problems. Nevertheless, shale oil and other ' synthetic fuels are expected to play an important role in new energy supplies in the longer term. WHAT IS OIL SHALE OR SHALE OIL? The term "oil shale" is sometimes a misnomer, in that the rock is often more of a limestone or siltstone than a shale. The common link between resources termed “oil shale" is that they all contain an insoluble substance cal led kerogen (which is from the Greek words for waxmaking). Kerogen is a form of organic carbon derived from a variety of plants ranging from algae to higher plants. When heated sufficiently, the kerogen generates hydrocarbons called shale oil, a form of synthetic crude oil that in most cases is lower in hydrogen content than conventional crude oil. The amount of oil in oil shale is relatively small --roughly 10 percent (by weight) in the richer shales. To upgrade this synthetic oil to usable products, additional processing is necessary. This brief sketch gives an idea of what this different, but significant, form of hydrocarbon is like. ENVIRONMENTS OF DEPOSITION Most oil shale deposits fall into three environments of sediment deposition: 1ake (called lacustrine), sea (marine) and river (fluvial-deltaic). In each case, the deposition of oil shales took place in quiet water environments where plant life, particularly algal plants, could flourish and, after dying, be deposited in unoxygenated water where the kerogen precursors would be safe from destruction by oxidation. The oil shales that were deposited in large lake basins (lacustrine) have attracted the most attention for development over the years. They often have multiple seams, deposited in a cyclic nature with extensive areal distribution and rapid vertical changes in kerogen content. Grades are moderate to high, ranging from 80 to 200 liters per tonne. Rundle in Australia and the Piceance Creek Basin in Colorado are examples of this type. Both deposits represent large volumes of oil shale in small areas which could provide the large volume of feedstock needed for future commercial operations. The stratigraphic sections of these two deposits feature thick oil shale seams with average grades of 80 - 125 liters/tonne conducive to both open pit, and underground operations. However, the rock strength of the Rundle shale is not sufficient to - support underground mining. On the other hand, the Colorado deposits, being more carbonate in nature, are sufficiently strong to support either type of mining depending on the overburden to ore ratio. These latter types of deposits will likely provide the first target for development of a commercial industry. The marine type is characterized by extensive areal distribution with relatively thin seams. The grades are generally low to moderate, ranging from 50 to 120 liters per tonne. The marine oil shales are common worldwide, and their attractiveness for mining is dependent on the overburden to ore ratio. Because of their widespread areal distribution, their in situ resources can be quite large. The Toolebuc Formation in Central , Queensland, Australia is a good example of this type of deposit being 7-10 meters thick over an extensive area. The Julia Creek deposit with its favorable overburden-to-ore ratio is being studied for possible development. In a fluvial-deltaic environment, there are many small lakes or bogs associated with rivers in which a very pure type of oil shale called torbanite could form. Torbanites are very high grade containing up to 75 percent hydrocarbons. The known occurrences are generally small lenticular deposits associated with coal seams. Even with the high grades, it is not likely that any of the known deposits would warrant commercial development because of their small size. The torbanite deposits in New South Wales, Australia were processed prior to World War I1 near the town of Glen Davis. However, today's known resources of this type are not large enough to warrant a commercial plant.

Jan 1, 1982

By J. J. Pye, J. B. Drew

Mzasurements of the surface diffusion coefficients of metals have been made. Diffusion profiles for the Ag-Ag system were obtained by means of a radioactive point source and a precision auto-radiographic technique. The activation energy for silver self diffusion (=8.1 kcal per mole) is lower than that previously reported (-10 kcal per mole) on poly crystalline wire by Nickerson and Parker. The bresent data indicate an effect due to parasitic volume diffusion at temperatures above 500°C. RELATIVELY few measurements have been made of the surface self-diffusion coefficients of metals. Nickerson and arker' measured the diffusion of silver over the surface of poly crystalline wires and estimated that the activation energy was 10.3 kcal per mole. Winegard and chalmers2 carried out measurements on both polycrystalline and single crystal surfaces but did not report a value of the activation energy. They found, however, that at temperatures between 250" and 400°C the diffusion coefficients were on the order of lo-' sq cm per sec and that there was an acceleration of the migration of silver on the polycrystalline sample when a change of surface shape occurred. Winegard and Chalmers used an autoradiographic technique, hereafter designated ARG, and Nickerson and Parker used a surface scanning geiger counter in order to determine the diffusion profiles. More recently, Hackerman and simpson3 measured the surface self-diffusion coefficient of copper at a single temperature (750°C), and the value of the diffusivity (- 10-5 sq cm per sec) is in agreement with that given by jostein from his thermal grooving measurements. This paper reports the results of an investigation of the surface self-diffusion coefficients of silver over a large temperature range and describes the adaptation of autoradiographic (ARG) techniques for the determination of diffusion profiles obtained from a radioactive point source. EXPERIMENTAL PROCEDURE The experimental procedure is a modification of the method employed by Hackerman and simpson3 in their measurements on copper. A brief description of their technique is as follows: A radioactive needle which sinters to the surface during the diffusion an- neal serves as the source of diffusing atoms. After the diffusion run the needle is removed and the surface is scanned with a shielded counting arrangement. The diffusion profiles reported in this paper were obtained by a modification of the above procedure which employs a precision ARG technique. Previous investigations in this laboratory and elsewhere51B have shown that under carefully controlled developing conditions and by the use of calibration sources a linear relation exists between the concentration of the isotope and the photographic density for values below unity. The use of ARG under these conditions has advantages over the counter scanning method in that cumbersome shielding and requirements for great mechanical precision of the scanner are eliminated. Also the ARG gives a complete picture of the surface which is advantageous in studies of anisotropic diffusion. A recording microdensitometer having a 0.1 p wide slit was employed. At low temperatures the disturbing effects of subsurface radiations are negligible. The diffusion anneals are carried out in the cell shown in Fig. 1. The needle is formed by grinding down a 1.0 mm rod of high-purity silver until a tip of 0.2 mm radius or smaller is formed. This tip is plated withA"' which becomes the source of the diffusing atoms that are detected by ARG. The needle carrier and the crystal holder, Fig. 1 are constructed of quartz and ports are provided in the holder pedestal which allow free vapor circulation ((2.0 oz) and the carrier apron fits snugly over the crystal holder cap, insuring that the needle does not move and scratch the surface. Temperatures are provided by a stabilized tubular furnace which can be quickly positioned around the cell, thus bringing the crystal up to temperature in a time that is short compared to the diffusing times. The diffusion anneals range from 2 hr for the high-temperature samples to about 25 hr for those at the lowest temperature. The possibility of vapor transport of the radioactive metal as a contributing factor in the diffusion profile was investigated in two ways. One method was to suspend the needle directly over a dummy sample, raise the temperature, for periods of time equal to the diffusion times, and then take an auto-radiograph of the surface. Negligible radioactivity appeared. In the second method a thin slot in the crystal face on one side of the source provided a "cong path" for surface diffusion. If evaporation was the primary source of surface atoms the region of radioactivity around the source would be symmetrical. This was not the case. The profile dipped abruptly at the edge of the slot but on the other side of the source the usual diffusion profile appeared.

Jan 1, 1963

By George Simkovich

It was hypothesized that solid-state point defects should alter the flotation properties of solids. Tests conducted on pure AgCl and AgCl doped with CdC12 show that atomic point defects exhibit an important role in the floatability of AgC1. Tests conducted on PbS doped with Ag2s or Bi2S3, also show that the defect structures resulting from these dope additions, i.e., a combination of electronic and atomic point defects, contribute significantly to the flotation of PbS. IT has been established that flotation occurs only when a finite contact angle exists between a solid and a gaseous bubble.' This angle, measured through the liquid phase, is expressed by the equation where the are inter facial free energies and the subscripts S, G, and L represent solid, gas, and liquid phases, respectively. As is seen in Eq. [I] three interface free energies, sG, sl, and GL, enter into the contact angle equation. Therefore, any variation in these energies which sufficiently varies the contact angle will, in turn, vary flotation processes. Changes made in any of the phases concerned, i.e., gas, liquid, or solid phase, are reflected through the changes occurring in two of the surface energy terms. Thus, a change in the liquid composition would be noted in sL and GL, and it is this phase, the liquid, which is most frequently altered in flotation studies., Changes in the solid phase must be reflected through the changes occurring in the sG and sL terms. In particular, it is hypothesized that changes in the surface concentrations of point defects in the solid-phase will alter the sG and sL terms which, in turn, will be reflected by flotation results. As an illustration of this hypothesis one may consider the defect structure and the flotation of AgC1. The bulk defect structure of AgCl is essentially one involving equal number of cation vacancies and interstitial cations.3 Upon adding CdC1, to AgC1, a greater number of silver ion vacancies are created in the bulk of the crystal.4 On the surface of the crystal the smaller binding forces and the free space accomodations may also allow for the creation of "surface interstitial anions", which will be designated as ad-anions. Thus, the point defect structure of the surface of AgCl doped with CdCl, will consist of cation vacancies and/or adanions. If the molecular forces responsible for the surface energies, ?SG and ?sL, are significantly altered by the presence of these surface point defects, then differences in flotation results will be noted as the concentration of these defects is varied. The defects present in AgCl are predominantly atomic in nature. In the case of PbS both electronic and atomic defects are present.5 This compound conducts electrically by either electrons or electron holes depending upon whether excess lead or excess sulfur is present. Upon disolving BiS3 in stoichio-metric PbS, one increases the concentration of cation vacancies and the number of electron carriers in the bulk of the crystal.5" At the surface, the possibility of ad-anions must also be considered. Conversely, upon dissolving AgS in stoichiometric PbS one increases the concentration of interstitial cations and the number of electronhole carriers in the bulk of the crystal.5,6' At the surface the interstitial cations will be designated as ad-cations. Thus, the point defect structure of the surface of a PbS crystal doped with Bi2S3 will consist of a number of cation vacancies and/or ad-anions and an excess of electrons. Conversely the point defects on the surface of a PbS crystal doped with Ag2S will consist of a number of ad-cations and an excess of electron holes. Again, as in the case of AgC1, should the molecular forces responsible for the magnitude of the interface free energies, ?sG and ?sL, be significantly altered by the presence of these surface defects then significant differences in flotation results will be noted as the concentration of these defects is varied. EXPERIMENTAL To test this hypothesis flotation tests were conducted on pure and doped AgCl and on PbS doped with either Bi2S3 or Ag2S. Preparation of the AgCl samples was performed as follows: AgCl and weighed amounts of CdC1, were melted in a porcelain crucible. The melt was then forced through a capillary tube and the particles emitted solidified in air as they fell about 1.5 meters. Spherical particles, -0.50 + 0.25 mm, were separated from the remaining solidified material

Jan 1, 1963

By T. S. Lovering

GEOCHEMICAL prospecting extends the age-old method of searching out lodes with a gold pan and rationalizes the prospector's hunch that certain plants are associated with ore. It uses sensitive but cheap and rapid analytical methods to find the diagnostic chemical variations related to hidden mineral deposits. Exploration geologists can gain tremendous assistance from this new tool, although its optimum use is not simple. To bring out the geochemical pattern that reveals the presence of a hidden ore deposit with a minimum number of samples requires a combination of shrewdness, chemical knowledge, and exploration geology. The use of sensitive analytical methods for prospecting had its start in the 1930's in northern Europe, where Scandinavian and Russian geologists had some success in these early efforts. Very little geochemical prospecting was carried on in the United States at this time, and no sustained interest was manifest until the close of World War 11, when geochemical investigations were started by the Mineral Deposits Branch of the U. S. Geological Survey. The purpose of these investigations was to apply geochemical principles and techniques to surface exploration for mineral deposits. Both the research on analytical methods and the routine trace analyses for field investigations were at first conducted by a single group, but it later became apparent that the trace analyses could be done by men of less experience than that required for successful research on methods. For the past several years there have been two groups of chemists, and although their functions overlap, three of the chemists are chiefly concerned with research, while four to six other men make the trace analyses for field projects. The chemical investigations, as well as the field projects of the Geochemical Exploration Section, concern only those phases of the subject that are appropriate to a government organization; every effort is made to help private industry, but not to compete with it, in finding orebodies. The chief aim of the Section, therefore, is to develop new analytical techniques and publish the results promptly, to carry out field investigations of the fundamental principles of geochemical dispersion, and to field test promising- techniques under controlled conditions. Some routine geochemical exploration work is carried on in connection with DMEA loans, and in district studies where the project chief wishes geochemical information on certain areas for his report. It should be emphasized, however, that geologists of the Geochemical Exploration Section are primarily concerned with fundamental principles underlying the distribution, migration, and concentration of elements in the earth's crust. To facilitate the use of geochemical methods the USGS has published much information on its methods of analysis and has provided opportunities from time to time for qualified professional personnel to study these methods, to work in the USGS laboratory, or to attend demonstrations of the analytical techniques at the Denver Federal Center. Typical of the research carried on are the problems now being investigated: 1) Development of rapid and sensitive analytical methods suitable to the determination of traces of metals and other minor elements in various materials, such as rock, soils, plants, and water. At the present time attention is being concentrated on U, Bi, Cr, and Hg, and satisfactory rapid trace analytical methods are virtually perfected for U and Bi. Good methods are also available for: Cu, Zn, Pb, Ni, Co, As, Sb, W, Mo, Ag, Nb, Ge, V, Ti, Fe, Mn, S, and P. 2) The relation of geochemical anomalies in plant materials to the geochemical distribution of elements in soils surrounding the plant. 3) A study of the dispersion halos in transported sedimentary cover such as glacial drift and alluvium over known orebodies. 4) A study of the behavior of ore metals in the weathering cycle. 5) A study of the behavior of the ore metals during magmatic differentiation. This requires a study of the distribution of minor metals in fresh igneous rocks and their component minerals in a well established differentiation series and in adjacent country rock. 6) A study of the dispersion of metals in primary halos in the wall rock surrounding orebodies. 7) Regional and local studies of the metal content of surface and groundwater in mineralized and barren areas. Many field projects of the Mineral Deposits Branch also require the services of USGS chemists during their investigation of the geochemical environment of ore deposits. From the work that has been done certain general principles have emerged. Concentrations of an element that are above the general or background value of barren material are called positive geochemical anomalies or simply an anomaly, whereas values less than background are called negative anomalies. The anomalies most commonly investigated in geochemical prospecting are those formed at the earth's surface by agents of weathering, erosion, or surficial transportation, but more and more attention is being given to primary anomalies found

Jan 1, 1956

A General Meeting and Colloquium was held in Kelvin House, Johannesburg on Wednesday 29th September, 1971. The theme of the Colloquium was "The Ferro-alloy Industry". Prof D. D. Howat (President) was in the Chair. The meeting was attended by 106 delegates and was opened by the President at 10.05 a.m. OBITUARY The President: "It is my sad duty to announce the death of G. C. Needham, Fellow, who joined the Institute in 1966 and passed away on 27th September, 1971. As a mark of respect to the memory of the deceased and in sympathy with the bereaved I would ask you all to rise and observe a few moments silence." MEMBERSHIP The President: "I have much pleasure in announcing that the names of the following candidates, having been published in accordance with By-Law 5.2.2, Council has elected them to membership of the Institute. Fellows: Richard Lloyd Blandy, Brian Roger Borton, Brian Teversham, Peter Donald Dickson. Members: Frederick Brian Peacock, Peter John Stacey, Kenneth Niles Thanstrom, Byron J. Hayley, Ian Flemming Buchan, Abraham Hercules le Roux. Graduate: Ian William Coy. Students: Peter Boesley J ames, Kenneth Brian Perel. TRANSFERS From Member to Fellow: Cecil Eugene Fivaz, Peter Edwards. From Student to Graduate: Roy Stavely Traviss. From Student to associate: William Mitchell. I welcome the new members to the Institute and congratulate those members who have been transferred to a higher grade. COLLOQUIUM The first paper, presented by Mr S. Selmer-Olsen, Consulting Metallurgist, Amcor Limited, on 'Trends in Ferro Alloy Production' was published in the May issue of the Journal and comprised information and data relating to new developments in ferro alloy production in various parts of the world. This was followed by a paper-'Conductivity and Viscosity Properties of Slags produced in ferro-alloy manufacture' by D. 1. Ossin et al. This was a record of laboratory test work on slags containing both lime and magnesia, corresponding with the composition of slags encountered in the production of ferro chrome alloys when a dolomite flux is employed. The third paper-'Review of the work of the Pyrometallurgical Research Group of N.I.M. and Metallurgy Department, Wits.' presented by Dr P. R. Jochens, leader of the Research Group, outlined the progress made on ten different research projects presently under way in the Department of Metallurgy. These projects range from studies of the kinetics and mechanism of the reduction of chromite ores through the determination of the activities of MnO in slags to phase studies and melting point determinations in the alloy systems Cr-Fe-Si-C and Cr-Fe-Si. A new feature of this Colloquium was the introduction of the discussion panel composed of four or five individuals known to have special knowledge or interest in the subject under duscussion. Mr J. J. Coetzee was chairman of the first panel, Professor D. D. Howat of the second and Dr R. E. Robinson of the third. During the discussion of Mr Selmer-Olsen's paper an interesting contribution was made by Mr Jack Hayley of Transalloys (Pty.) Limited who pointed out that the lowest power cost in the U.S.A. was that from the TV A at 6 mils per unit (0,43 cents S.A.) while costs from other sources were 7 to 8 mils per unit (0,50 to 0,57 cents S.A.) He also referred to the use of 'pile sintering' of ore fines plus coke at certain ferro alloy plants in the V.S. and to the recent practice of 'graphitising' the paste used for Soderberg electrodes. The last session in the afternoon was devoted to a group discussion on 'Practical problems in Arc Furnace Operation', Mr H. R. Reid of Middelburg Steel Alloys (Pty) Limited occupying the chair, members of the panel being Mr J. Meintjies (Rand Carbide Limited), Mr J. Benn (Amcor) and Mr E. J. Jackson (Transalloys (Pty.) Limited). Mr Reid initiated the discussion by talking of the design and composition of electrode shoes, mentioning the use of the 85/15 Cu/Zn alloy and the comparison of hollow shoes and shoes cooled by tubes. Mr Meintjies discussed electrode management with particular reference to breakages, and dealt with casing and fin design and the welding of fins. Mr Benn dealt with aspects of closed top furnace operation with particular reference to the control of eruptions in the furnace and indicated that preheating of the raw materials might be well worth considering together with the development of a system for the detection of the water vapour present. Mr J ackson was concerned with shells, linings, refractories and tap holes. He deprecated the use of oxygen for opening the tap hole emphasising how uncontrolled use of oxygen led to loss of shape of the tap hole and quite severe burning out of the front hole resulting in time and labour in repairs. A notable feature of the Colloquium was the free discussion 'from the floor' of the hall. Many valuable and highly pertinent contributions were made spontaneously and usually in reply to some controversial topic raised by the discussion group.

By A. J. Shaler, H. Udin, J. Wulff

In the study of the sintering of meta powders, we have come to the conclusion in this laboratory that further progress requires a more basic understanding of the operating mechanisms. This is emphasized in detail by Shaler. He has shown that a knowledge of the exact value of the surface tension is imperative for a solution of the kinetics of sintering. This force plays a principal role in causing the density of compacts to increase.2 Furthermore, a knowledge of the surface tension of solids is also applicable to other aspects of physical metallurgy. C. S. Smith3 points out the relation between surface and interfacial tension and their function in determining the microstructure and resulting properties of polycrystal-line and polyphase alloys. This paper describes one group of results of an experimental program designed for the study of the surface tension in solid metals. As a by-product of this work, considerable information has been obtained on the rate and nature of the flow of a metal at temperatures approaching the melting point and under extremely low stresses, a field of mechanical behavior heretofore scarcely touched by metallurgists. The importance of this additional information to students of powder metallurgy need not be stressed. Theoretical Considerations Interfacial tension arises from the condition that an excess of energy exists at the interface between two phases. Gibbs proves that this energy is a partial function of the interfacial area; thus: ?F/?s = ? where ?F/?s is the rate of change of free energy of the system with changing surface area, at constant temperature, pressure and composition, and ? is the interfacial tension, or interfacial free energy per unit area. If one of the phases is the pure liquid or solid, and the other the vapor of the substance, ? may properly be termed "surface tension," and is a characteristic of the solid or liquid. The attempt of a body to lower its free energy by decreasing its surface gives rise to a force in the surface which is numerically equal in terms of unit length to the free energy per unit area of the surface. Thus ? may be expressed either in erg-cm-² or in dyne-cm-1. Similarly, surface tension may be determined either by a thermo-dynamic measurement of the surface energy or by a mechanical measurement of the surface force. We have chosen the latter approach. Tammann and Boehme4 determined the surface tension of gold by measuring the amount of shrinkage or extension of thin weighted foil at various temperatures and interpolating to zero strain. The method is of questionable accuracy because of the tendency of foil to form minute tears when heated under tension. Their assumption of F = 2W?, where W is the width of the foil, is unsound, as the foil can decrease its surface area by transverse as well as by longitudinal shrinkage. Although their experimentation was meticulous, the paper does not include details of the sample configuration required for recalculating ? on a correct basis, even if such a calculation were possible. In the experimental procedure chosen here, a series of small weights of increasing magnitude are suspended from a series of line copper wires of uniform cross-section. This array is brought to a temperature at which creep is appreciable under extremely small stress. If the weight overbalances the contracting force of surface tension, the wire stretches; otherwise, it shrinks. The magnitude of the strain is determined by the amount of unbalance, so a plot of strain vs. load should cross the zero strain axis at w = F?. If balance is visualized as a thermodynamic equilibrium, the critical load is readily calculated. At constant temperature, an infinitesimal change in surface energy should be equal to the work done on or by the weight: ds = wdl [A] For a cylinder, s = 2pr2 + 2prl [2] If the volume remains constant, r = vV/pl [31 s = 2vpl+2V/l [4] ds = vpv/l - 2V/l²) dl [5] Substituting [5] into [I] gives for the equilibrium load, w = ?(z/rV- 2V/12) [6] and, again expressing V in terms of r and l, w = pr?(1 - 2r/l [7] Here the end-effect term, 2r/l, is neglected for thin wires in subsequent work. Eq 7 can be confirmed by means of a stress analysis. If the x-axis is chosen along the wire, then the stress is 2pr? - w pr² pr2 [8] A cylinder of diameter dis equivalent to a sphere of radius r, insofar as radial surface tension effects are concerned.³ Thus xv = 2?/d = ?/r = sz [9] For the case of zero strain in the x direction, the strain will also be zero in the y and z directions. Since the wire is under hydrostatic stress, Eq 8 and 9 are

Jan 1, 1950

By T. Cantwell, N. C. Rasmussen, H. E. Hawkes

Beryl is a mineral that may be difficult to distinguish from quartz by casual field inspection. The easily recognized green color and hexagonal crystal form of coarse-grained beryl are by no means universal, even in beryl from pegmatitic deposits. If it occurred as a fine-grained accessory mineral in an igneous rock, it would almost certainly escape detection unless samples were submitted for petrographic or chemical analysis. There may be substantial deposits of some beryllium mineral, other than beryl, that has been overlooked because that mineral also closely resembles the common rock-forming minerals. A reliable and simple method of identifying beryllium minerals and determining the beryllium content of a rock would be helpful in exploration. This article describes preliminary experiments in applying nuclear reaction to the qualitative identification of beryl and to the semiquantitative determination of the beryllium content of rock samples. Gaudin,1,2 the first to apply a nuclear reaction in detecting beryllium minerals, developed a method that irradiates the sample with gamma rays, which react with beryllium nuclei to produce neutrons. The neutrons are then measured with standard equipment. The cross section for this reaction is about 1 millibarn. The cross section is a measure of the probability that a reaction will take place, for example, between a beryllium nucleus and an incident gamma ray or alpha particles.3-5 At 1-millibarn cross section for the reaction, satisfactory performance required a source strength of the order of 1 curie (3.7 x 10"' disintegrations per sec, where each disintegration releases one or more gamma rays). The reactions will not take place if the gamma radiation is below a minimum energy, in this case 1.63 mev. The size of the source and the energy of the radiation made heavy shielding necessary for these experiments, both to reduce the background count of the neutron counter and to safeguard personnel. The original discovery of the neutron by Chad-wick in 1932 resulted from experiments with another nuclear reaction, induced by bombarding beryllium with alpha particles in which the products are carbon-12 and neutrons. The equation for this reaction is as follows:' " ,Be" + ,He'? 6C12 + 8,n' [1] re-particle neutron In the above nuclear equation (Eq. 1), the sub- script number indicates the number of protons in the nucleus (the atomic number) and the superscript the total number of neutrons and protons (approximately the atomic mass). For the alpha-neutron reaction the cross section is about 250 milli-barns, or 250 times that of the gamma-neutron reaction used by Gaudin. The positively charged alpha particle is repelled by the positive charge of the beryllium nucleus; it must, therefore, have a certain minimum energy in order to approach close enough to the beryllium nucleus to react. For reaction with the beryllium nucleus, the lower limit of the alpha-particle energy is 3.7 mev. The alpha-neutron reaction, with polonium-210 as an alpha source, was selected for the present experiments. Alpha particles are emitted by polonium-210 at 5.30 mev, which is adequate for the reaction with beryllium. Furthermore, this isotope of polonium emits alpha particles with negligible associated gamma radiation, thus eliminating the necessity of shielding. The half-life of polonium-210 is 138 days. Inasmuch as alpha particles carry a possible charge and are large compared with most nuclear particles, their energy is rapidly dissipated in passing through matter. Their range in standard air is 3.66 cm,3 and they penetrate only a few tens of microns into a mineral sample. The short range in air can be minimized by preparation of a flat sample surface that can be brought very close to the alpha source during analysis. On the other hand, short range of alpha particles in air lessens the radiological health hazard and makes it possible to use this method without shielding. It must be emphasized, however, that the alpha emitters are potentially very dangerous if they enter the human body. Polonium must be handled with extreme caution. The literature has reported experiments on the yield of neutrons from reaction of alpha particles with beryllium nuclei. Feld" reports that in intimate mixtures of polonium and beryllium, 3 x 106 eutrons per sec are produced per curie of polonium. Elsewhere in the same reference it is stated that a sandwich-type source yields about one third as many neutrons as an intimate mixture. A table of neutron yields for full energy polonium alpha-particles on thick targets as reported by Anderson7 is the basis of Table I. From Table I it can be deduced that the elements most likely to interfere, i.e., those that also produce neutrons when bombarded by alpha particles, are boron and fluorine. These data also show that it will probably not be possible to determine very small quantities of beryllium in rocks because of the masking effects of the major elements, sodium, magnesium, and aluminum. The neutrons emitted in the alpha reaction are detected by another nuclear reaction. Either of the

Jan 1, 1960

By A. J. Richardson

Before the development of the underground stoping and mining costs can be considered, certain facts about the ore body, the proposed mine, markets, etc., must be known or determined. In the case to be studied, the zinc-lead mineralization occurred with a narrow vertically dipping structure of undetermined length and vertical extent. Exploration completed to date has revealed 6.5 mil¬lion st t of proven reserves. A further 820,000 st of in¬dicated reserves has been outlined and this tonnage is considered capable of being expanded by a factor of approximately four after more detailed drilling. After studying the market conditions and completing a very preliminary feasibility study, it was decided that production would be 730,000 stpy (or 2000 stpd) of ore. First year production would be at the rate of 1500 stpd. The main design criteria for the selection of the min¬ing methods are minimizing surface subsidence, maxi¬mum recovery of the ore body, maximum degree of grade control, maximum productivity, and safe working conditions. Two basic extraction systems are considered capable of meeting these requirements: mechanized cut¬and-fill stoping and sublevel long-hole stoping with filling. The primary development system of the mine has been designed to give maximum flexibility in stoping systems and layout and to permit changes if considered necessary as a consequence of actual production ex¬perience. At the present time, access to the mine is by a circu¬lar concrete lined vertical shaft, 16 ft diam, sunk to a depth of 1380 ft. Two exploration levels have been driven within the ore zone at depths of 165 and 1246 ft below the surface outcrop. The development to date had the objective of sampling the mineralization and produc¬ing detailed information on the outline of the ore body and the distribution and controls of zinc and lead values. In an attempt to satisfy the basic design criteria for the mine, it was decided that production would be best achieved by a combination of 40% sublevel long-hole stoping and 60% cut-and-fill mining. Costs of exploration and capital development of per¬manent underground facilities are normally written off over the life of a mine. Production expenditures, on the other hand, are of a temporary nature and are normally charged as and when incurred as an operating expense. Reasonably accurate predictions of mine production costs can be built up from engineering design and estimates of individual mine activities for ultimate inclusion in the comprehen¬sive data required for financial decision making. The simulated operations can be costed on a detailed basis in the form of a monthly operating budget. The budget format can be generalized or detailed, depending upon the scope of the project. However, ex¬perience suggests that a fairly detailed format has the advantage of assuring that all significant cost items are included. For underground costing it is suggested that the budget structure include five major cost centers (i.e. development, diamond drilling, ore extraction, hoisting/ transportation, and general mine expense). These in turn are detailed under numerous subheadings. The mechanism for compiling an operating budget will be illustrated. Because of its relative simplicity, ore extraction under sublevel long-hole stoping has been chosen for illustration. All other activities, simple or complex, can be estimated in similar fashion. BLOCK AND STOPE DEVELOPMENT Long-hole blocks, used where advantageous, will be up to 250 ft in height, depending upon the vertical con¬tinuity of the mineralization, and approximately 300 ft long. Drawpoints will be at 36-ft intervals and serviced by loading crosscuts driven from a footwall drift parallel to and close to the ore zone. Pillars between the stopes will be 50 ft wide. Stopes will be drilled off with vertical rings of blastholes drilled from sublevels approximately 60 ft apart vertically. This drilling will be done by percussion drilling machines (31/2 in.) mounted on a trackless drilling rig. Load¬haul-dump (LHD) equipment will be used to move broken ore from the drawpoints to the orepass connecting to rail haulage systems. On completion, long-hole stopes will be backfilled to prevent caving and to facili¬tate later pillar removal. From a planned stope layout, a forecast of produc¬tion and development is made in Table 1. Table 1. Block Tonnage and Stope Development Quantity Ore Waste Total ore block 375,000 st 2 stopes 310,000 st 1 pillar 65,000 st Access crosscuts, 4 at 100 ft 400 ft Drill sublevel drifts, 6 at 300 ft 1800 ft Stope raises, 3 at 250 ft 750 ft Undercut sublevel drifts, 2 at 300 ft 600 ft Loadout crosscuts at 35-ft intervals 550 ft 100 ft 3300 ft 500 ft Total development footage 3800 ft Tons per ft of development 987 st

Jan 1, 1982

By R. Masse, J. Chameaud, J. Lafuma, R. Perraud

Estimating the risk of lung cancers for workers in uranium mines and defining the resulting dose equivalent limits have been made possible thanks to work carried out in two scientific fields : physics and epidemiology. Theoretical calculations on the basis of physical models for the former and epidemiological surveys on the mortality of uranium miners from lung cancer for the latter. However, even though considerable work has been done in these two areas, the results obtained still remain controversial on several points. The radioactive and physical instability of the aerosols present in the atmosphere of the mines and the biological complexity of the human lung which even the most sophisticated physical models can reproduce only very schematically have often proved to be insurmountable difficulties for physicists, explaining the uncertainties which subsist concerning the dose delivered to the various parts of the respiratory tract from the air breathed by miners during their work. Epidemiological investigations on the other hand, in spite of the high quality of the surveys carried out, remain open to criticism, essentially because of the very approximative estimation of the individual occupational exposure to radon daughters. This is due to the fact that uncertainties arise from the measurement of radon gas if the state of equilibrium with the daughters is not accurately known or, if the active deposit is measured, to the fact that these measurements are insufficient in number. The controversies and discrepancies which subsist with regard to the evaluation of the level of risk, and in particular for low doses, can thus be understood. In addition, epidemiological surveys cannot dissociate the carcinogenic action of radon from the synergistic or potentiating actions of tabacco and of other pollutants present in uranium mine air. Animal experiments have been largely taken into account for evaluating the toxicity of various radionuclides. This type of experiment is necessary when human data do not exist and has provided us with much information. For instance, the relative biological effectiveness of the various types of radiation, the metabolism of radionuclides and the mechanisms of cancer induction have been approached and satisfactorily resolved in this way. Concerning radon and its daughters, however, animal experiments have been used very little even though it seems apparent that they should complement epidemiological studies. For instance, whereas doubt can be cast on the data obtained from human epidemiology because of the uncertainty concerning the individual exposure of miners, those drawn from experiments are indisputable because in this case the dose is as perfectly known as the effect. In addition, the effects of radon can experimentally be appreciated separately whereas in the surveys, they cannot be dissociated from the effects of the other pollutants in the mine. Finally, there are no other means of dealing with the mechanisms of cancer induction. In order to gain any useful knowledge from this method however, the experimental model must necessarily present certain methodological guarantees and the effects seen in the animals must enable a comparison with those which appear in man. For this reason we will present here the animal model we have been using for 15 years, and will give the results obtained and compare them with human data and made a synthesis. Finally the conclusions which can be drawn will be discussed as well as their limitations with respect to the protection of uranium miners. I - MATERIAL AND METHODS Male SPF Sprague-Dawley rats were used. At the onset of the inhalations they were around 3 months old. Their small size makes it possible to expose a large number of animals at the same time. Their life-span is long enough to be able to follow the evolution of the cancers and to estimate the latency time. Finally, they present the advantage of having a very low rate of spontaneous lung cancers (SANDERS, 1979). I.1 - Three inhalation techniques were used. 1.1.1. – [Inhalation of radon decay products.] The inhalation apparatus has been described previously (CHAMEAUD et al. 1971). The first experiments utilized a room of a half cubic meter linked to a source made up of high grade uranium ore. Later on, a large installation was built with a 10 m3 inhalation chamber making it possible to expose up to 500 rats at one time at radon concentrations ranging from 100 to 10 000 WL for variable lengths of time (1 to 10 hours per day). These concentrations are higher than those to which the miners are generally exposed, but in order for the cumulated doses in man and in animal to be similar and delivered for the same fraction of their respective life-spans, the ratio of the concentrations should be approximately that of the life-spans. The concentrations of radon and its daughters during the experiments were carefully controlled thanks to multiple samplings of radon gas associated with measurements of radon decay products. I.1.2 – [ The dust inhalation chamber] has already been described : it is a dust-loading chamber where the dust content remains constant during the experiment and can hold over 20 - 30 animals (PERRAUD et al, 1970). I.1.3 – [Tobacco inhalations] take place in a smoke box

Jan 1, 1981

By Larry D. Patts

Section 317(e) of the Federal Coal Mine Health & Safety Act of 1969 directed the Secretary of the Interior to prepare standards under which all working places in a mine shall be illuminated by permissible lighting while persons are working in such places. Section 317(e) further provides that such working places shall be illuminated within 18 months after such standards are promulgated. In accordance with this section of the Act, there was published in the Federal Register for December 91, 1970, a notice of proposed rulemaking which prescribed the illumination to be provided in the working places of underground coal mines. In light of written comments, suggestions, and objections to this proposed rulemaking, the proposedstandards were withdrawn and reproposed in the Federal Register for Wednesday, October 27,-19h. In light of further comments, suggestions, and objections, a public hearing was held on April 4, 1974, and standards were again reproposed and published in the Federal Register for April 1, 1976. Promulgation of the final lighting standards took place on October 1, 1976, which means that the underground coal mining industry must comply with face illumination requirements by April 1, 1978. As mentioned previously, the first proposed rulemaking for illumination of underground coal mines was published in the Federal Register on October 27. 1971. In early 1972, Consolidation Coal Company (Consol) and the United States Bureau of Mines agreed to a cooperative study of underground face illumination: Consol felt that expertise is this field would become increasingly important. Consol's initial efforts in illumination were aimed at investigating practical lighting systems for underground face equipment. We were concerned with installing unobtrusive lights which provided sufficient face illumination for safety, but at the same time were readily maintainable, electrically reliable, and physically sheltered from damage. We believe that our initial lighting systems did provide sufficient face lighting for safety, but because only prototype components were available for field testing, the resultant poor system reliability and maintainability necessitated drastic improvement before face lighting could become practical. Final Lighting Standards Deem Early Lighting Installations Out Of Compliance On April 1, 1976, the Federal Register contained the final version of the illumination standards (as they were later promulgated in October). When these illumination regulations and measurement techniques were defined and measuring instruments were available, Consol checked their lighting systems underground and determined that the systems were not in compliance with these final illumination standards. More Lighting Hardware Added In An Attempt To Achieve Compliance. After determining that all of our face lighting systems were not in compliance, we began adding additional lighting hardware in order to meet compliance with published regulations. Unfortunately, to date, we have not been able to meet compliance with practical lighting systems. We have determined from our field installations that the required additional lighting hardware, (to meet compliance) with its increased vulnerability and decreased reliability, renders the lighting systems impractical, if not impossible, to reasonably maintain. Our attempts to provide 0.06 footlamberts have also produced adverse operator reaction to the glare and to illumination systems in general. BCOA Members Ask MESA To Demonstrate Practicability Of Compliance With Regulations Industry concern about meeting the impending lighting regulations was mounting, and in May of 1976 a meeting between MESA and BCOA members was held to discuss lighting compliance problems. At this meeting, BCOA offered to work cooperatively with MESA in testing the practicability of various lighting systems mounted on underground mining equipment. Field tests were to be conducted by United States Steel Corporation, American Electric Power Service Corporation, and Consolidation Coal Company. The purpose of this underground field testing was to develop capability to provide adequate face illumination in a safe, workable manner which would not detract from efficiency of operation. BCOA members involved in this cooperative study were to submit necessary machine drawings, sketches, etc. to MESA in order that MESA could perform a "black-box" study and specify the type and location of luminaires to be installed on the test machines. MESA was confident that they could specify lighting systems that would be in compliance and would be practical so as not to detract from efficiency of operation. Consol was first to install lighting hardware under the BCOA/MESA cooperative agreement. As per MESA specifications, Control Products HgV luminaires were installed on a Joy 2BT-2H boring machine at the Williams Mine of Northern West Virginia Region. As of mid-October, 1976, Consol had approximately eight weeks operating experience with the lighting system on this boring machine underground and had drawn the following conclusions: The lighting system installed at Williams Mine (1) does not meet compliance with lighting standards as originally proposed by MESA, (2) does not provide illumination in a safe workable manner, and (3) will detract from efficiency of the mining operation due to operational delays. Although Consol has rearranged lights on this boring machine in an attempt to reduce operator objections, a practical lighting system which is "in compliance" has not been arrived at as of this writing.

Jan 1, 1979

By K. E. Barrett

Exxon owns a uranium mill and holds two mining leases in Live Oak County, Texas, about halfway between San Antonio and Corpus Christi. The properties make up the Felder Uranium Operations which was reopened earlier this year. Exxon held an oil, gas, and other minerals lease on the J. C. Felder tract, which was adjacent to a relatively shallow uranium discovery by Susquehanna-Western, Inc. on the Marrs-McLean lease immediately south of the Felder property. Drilling in 1967 and 1968 confirmed the presence of reduced uranium mineralization in the basal sand unit of the Oakville formation on the Felder tract, which formed the major part of the roll-front deposit. In 1969 Exxon and Susquehanna-Western, Inc. entered into a sale and purchase agreement which provided for Susquehanna to mine and process Felder ore and purchase recovered uranium. Susquehanna moved an alkaline-leach mill from Wyoming, erected it on the Ray Point property, and placed it into operation late in 1970. Susquehanna mined and processed ore from the Felder and McLean properties through March 1973. Susquehanna ceased operations in March 1973. Exxon then acquired the mill and mill property. Exxon also purchased the mineral rights to the McLean lease, re-negotiated a mining lease for that property, and carried out shut-down programs for the mining and mill areas in the fall of 1973. The project was put on a standby basis until late 1973, when Exxon initiated mine feasibility studies for the project. MINE PLANNING EVALUATION The feasibility study for reopening the Felder Project began in late 1975 and was not completed until late 1976. I will discuss several areas of the feasibility study that required additional work prior to making the decision to renew operations. Ore Reserves Preparations for estimating the ore reserves began with the re-evaluation of more than 1500 natural radioactivity logs from exploration and pre-development drilling that had been completed on the property. These gamma ray logs of non-core rotary drill holes were the principal source of data used in making the estimate. Chemical assays of cores from the deposit were also used in the reserve determination. Electric resistivity and self-potential logs were made along with the gamma ray log. In December 1975 an additional core drilling project was undertaken to confirm the in-place density and radiometric equilibrium characteristics of the ore deposits. Comparison of chemical assays of cores with the U308 values calculated from the logs showed that the unoxidized ores were in radiometric equilibrium. In contrast, cores from anomalies occurring in near surface, weathered, and oxidized zones were in radiometric disequilibrium. Several important decisions were made in developing a mine plan or schedule of production from the Felder and McLean ore bodies. Disposal of Produced Mine Water: The ore bodies of the Felder Uranium Project occur at a point below the ground water table. The ore zones to be mined must first be dewatered to allow removal of mineralized material. In the open pit operations, this is accomplished by maintaining a perimeter ditch around the periphery of the open pit, allowing the interior of the pit to drain and collect into a sump and be pumped from the mine. In addition to anticipated water production from future mining operations, approximately 200M gallons of water was contained in three open pits left from prior mining operations. In two of these existing pits, the water was to be removed and disposed to allow for planned backfilling of waste material into these pits. The third pit would also have to be drained to allow continued mining of an area left from the prior operations. Essentially no ground water information was available for this area. The only data available was water production history from Susquehanna's mining operation. Two water wells were drilled early in 1976 on the Felder lease for use in obtaining hydrological data. A long term draw-down test was performed by pumping one water well and measuring water level drawdown in both the pumped well and the observation well. From these data, values for permeability and storage coefficient were calculated. These data were then used in modeling the aquifer to allow calculation of water influx into the mining area versus time. Once a schedule of water production, including the stored volume in the existing pits was calculated, alternate solutions for disposal were evaluated. The first system evaluated was a series of deep injection wells. The wells were designed to inject at a depth of approximately 3500 feet. Again very little information concerning reservoir characteristics of the receiving sand units was known. Using assumed values for reservoir permeability and storage coefficients, an injection well system was designed to allow for disposal of produced mine water. The biggest

Jan 1, 1979

By R. G. Davies

The activation energy for steady state creep above -500°C is observed to be independent of the applied stress although it varies from -67 kcal per mole at 2 at. pct Si to -100 kcal per mole at 11 at. pct Si due to changes in crystallographic order. The magnitude of the activation energy, by comparison with Fe-A1 alloys, indicates FeSi type of order in certain alloys. X-ray results confirmed the presence of FeSi type of order. It is proposed that dislocation climb is the rate controlling mechanism for all the alloys. It has been demonstrated that when a diffusion mechanism is the rate controlling process, the formation of a superlattice in brass,1 Fe3A1,2 Ni3Fe,3-5 and Feco6 1) increases the creep resistance, and 2) increases the activation energy for steady state creep. Furthermore, a study of creep in Fe-15 to 20 at. pct A1 alloys7 has revealed that as the alloy composition approaches the long-range order field, there is an increase in the activation energy for steady state creep which is thought to be due to an increase in short range order. Fe-A1 and Fe-Si alloys are similar in that they both form the DO3 superlattice in which aluminum or silicon atoms have only iron atoms as first and second nearest neighbors. There are, however, two important differences between the alloy systems: 1) The superlattice formation at -350°C commences at -10 at. pct si8 as compared to -20 at. pct Al,9 and 2) Fe-A1 alloys form a FeAl (B2 type) super-lattice where aluminum atoms have all iron first nearest neighbors even at 22 at. pct Al, but so far no similar FeSi superlattice has been observed. With the similarity between Fe-A1 and Fe-Si alloys in mind, alloys of iron with 2 to 11 at. pct Si were examined for variations with composition of the activation energy for steady state creep and of creep strength. The temperature range of greatest interest was above 1/2 TM (TM is the absolute melting temperature) where it is usually observed that diffusion is the rate controlling process. A subsidiary X-ray investigation of the Fe-Si system was undertaken in an attempt to define the position of the order-disorder boundary as a function of cooling rate. EXPERIMENTAL DETAILS a) Creep. Specimens whose gage length was 1.5 in. and with a cross-section 0.04 by 0.08 in. were strained in tension by a lever-arm arrangement, and the load was adjusted between each creep test to maintain constant stress. The apparatus and mode of operation have been fully described in a previous publication.7 As each test produced a creep strain of 0.25 pct, the variation in stress during the test was negligible. Creep strain was measured at the end of one of the alloy steel grips by a displacement transducer with the out-of-balance potential being recorded on a variable speed recorder. The full-scale deflection of the recorder could be varied in steps to give limits of sensitivity of between 0.1 and 0.001 pct creep strain. The alloys, Table I, were made available by the Metallurgical Department, National Physical Laboratory (N.P.L.), england,10 and by the Research Department, General Electric Co. (G.E.), Schenectady, N.Y. They were hot worked at -850°C, warm worked at 550° to 650°C, and recrystallized in vacuum at -750°C to give a grain diameter of -0.1 mm. All the alloys had a very low impurity content; those from the N.P.L., for which a complete analysis is available,&apos;&apos; show carbon less than 0.026 pct, manganese less than 0.006 pct, and oxygen plus nitrogen less than 0.0024 pct. b) X-ray Procedure. A General Electric XRD-5 X-ray set with a focussing lithium fluoride mono-chromator in the diffracted beam, and a pulse height analyzer to eliminate harmonic wavelengths of the cobalt radiation, was used to investigate the structure of several very fine grained (grain diameter <.01 mm) Fe-Si alloys after the following heat treatments: 1) Quenched from 700°C, 2) slow cooled from 650°C (-40°C per hr), and 3) very slowly cooled from 400° to 100°C (10°C per hr with a 24 hr anneal every 100°C). The method of obtaining the diffraction pattern over the range of 20 from 15 to 45 deg was to count for at least 100 sec every l/3 deg with a slit subtending 1 deg in 20 at the focus; the probable counting error was less than 2 pct.

Jan 1, 1963

By Paul R. Bluekamp

The Mather Mine property is composed of a 5.2 sq km (2 sq mile) area within the Cities of Ishpeming and Negaunee which are located in the Upper Peninsula of Michigan. Production in this mine started in 1943; it ran continuously until its closing in 1979, having produced slightly over 55 million tons of iron ore. This mine was a joint venture between several steel companies and The Cleveland- Cliffs Iron Company (CCI), with CCI being the fee holder and operator. For the first 17 years of operation, the ore was shipped in its natural state directly from the mine to the steel mill. By 1960, iron ore pellets were on the market and they proved so superior to the natural soft Mather ore that the latter became difficult to sell except for the coarser portions. It was decided to develop a pelletizing process for the Mather ore, and this was accomplished by 1965 when the first Mather pellets were produced. From this date, all but the coarse fraction of the Mather ore was shipped as pellets. The geological setting is that of a large east-west trending synclinorium which plunges to the west. The lowest member of this trough-like structure is, for the most part, a quartzite-like graywacke, the upper 20 m of which grades into a softer, fine grained slate. Lying conformably on this graywacke footwall member is an iron- formation member which is over 1,300 m in thickness. This iron-formation is cow posed of thin alternate bands of iron oxides and chert and is intruded by a number of diorite sills - some up to 122 rn (400 ft) in thickness. The north limb of this synclinorium dips at approximately 45&apos; and bottoms out at about 1,000 m from surface in the central part of the mine. There are two sets of faults, many of which are intruded by diorite dikes, which trend east-west and southeast-northwest. Displacements are varied, reaching a maximum displacement of 243 m (800 ft). The ore is found lying directly on the slatey footwall and its position is largely controlled by the faults and dikes, with the bulk of the ore being on the upper side of these structures. The ore is composed of soft earthy hematite and martite with vertical thicknesses up to 122 m (400 ft) although the average thickness would be closer to 46 m (150 ft). The ore averaged 60% iron and 7% silica on a dried analysis. The Mather Mine is located on the north limb of the syncline and was worked from two shafts, the deepest of which. was 1,09 7 m. These two shafts are about 2 km apart and serviced a total of 8 working levels between them during the life of the mine. The level spacing was about 61 m (200 ft). The main haulageways were driven parallel to the ore/footwall contact in the hard cow petent graywacke wherever possible. On the lower levels, deeper into the footwall, naked development was common. This material graded into roof bolting ground towards the upper stratigraphic portion. As drifting progressed further into the upper stratigraphic portion of the footwall, progressively stronger steel sets had to be used. From the main haulageway, cross-cuts were turned into the orebody on 61 m (200 ft) centers and extended as far as needed to recover the ore available to that particular cross-cut. The main haulageways and cross-cuts were driven 3 m (10 ft) high, 3 m (10 ft) wide at the top and 4.6 m (15 ft) wide at the bottom. This configuration would accomodate a set composed of a 2.7 m (9 ft) cap on top of two 2.7 m (9 ft) legs angled out at 18&apos;. A sill plate was used under each leg to prevent its sinking into the ore below. The type of steel used was dependent on the expected weight to be experienced. Sets were placed on 1.63 m (5 ft 4 in.) centers; however, in extremely heavy areas, it was sometimes necessary to install sets on 0.8 m (2 ft 8 in.) centers. On the top two levels (5th and 6th), the ore was considerably hard-er and was mined by sub-level stoping and long hole drilling. Very light steel sets were used in the main haulageways and cross-cuts and timber was used in the production drifts. Some con- crete production drifts were installed on 6th level, but proved to be uneconomical. However, as mining reached greater depths, the ore became softer and more massive, reaching its maximum vertical heights on 11th and 12th levels. On levels 7 through 10, yielding steel sets were used extensively in the slusher drifts. While they were satisfactory on 7th and 8th levels, their success diminished with depth and they were

Jan 1, 1981

By A. L. Labbe

SOON after the Cottrell treater was placed in operation in the Murray plant in 1918 to treat combined lead sinter and Wedge roaster smoke, it was noticed that the power flowing through the treater did not remain constant. This was indicated by the varying milliamperes and also by the total amount of power consumed by the rectifiers. At times, for then unknown reasons, the treater current fluctuated through a wide range of from 40 to 300 milliamperes. Fortunately, these variations in power did not affect the treater&apos;s overall recovery, as this installation consisted of three independent units in series, a feature which made the Murray Cottrell an outstanding installation over many years of operation. Water conditioning of the smoke as a means of improving recovery was already known, but was not adaptable to this installation for reasons of excessive corrosion, which had been the case with other treaters using water as a conditioner exclusively. Tests conducted on the smoke had proved conclusively that water vapor played no part in the fluctuation of power, and the same was true of the SO2 contents and dust burden of the smoke. Neither did temperature variations of from 150 to 400°F have any effect on the power. Finally it was noticed that the variation in power taken by the treater, and the efficiency of recovery, had some definite relation to the number of Wedge roasters operating, and particularly to the sulphur contents of the charge. This observation soon led to the discovery that free sulphuric acid was the conditioner for Murray smoke and that variations in acid contents of smoke accounted for fluctuations of treater power. Analysis of recovered dust revealed that only a few hundredths of a per cent free acid was necessary to maintain a very efficient recovery. Once this knowledge was available, the roaster charge was adjusted as to sulphur contents to produce the necessary acid conditioner. For a number of years this practice was followed, but with improvements in the field of flotation, excess pyrites were eliminated from the smelting picture, so with a change in metallurgical practice we were confronted with the problem of providing the deficiency in acid by some other means. The situation was aggravated and our problem of acid conditioning made more difficult by the increase in the lead contents of concentrates roasted resulting from better flotation methods. The first step towards introducing acid vapors into the smoke stream by accessory means was accomplished by boiling sulphuric acid in cast iron pots placed in an open fire box. Acid fumes evaporated from the pots together with the combustion gases from the fire box were discharged into the flue through a cast iron pipe. Capacity of the iron pots was quite limited because of the comparatively slow rate of evaporation and the use of lump coal as a fuel, which did not lend itself to practical temperature control. This method of firing resulted in frequent pot failures due to overheating. In spite of these incidental difficulties, encountered in any new venture, the pot evaporator demonstrated the practicability of fuming sufficient acid to make up the deficiency in that naturally evolved in the Wedge furnace operation. As time progressed, less and less high sulphur material was available for Wedge roasting, and the necessity for accessory acid conditioning reached a climax when the supply of this material was entirely eliminated. Since the cast iron pots could only evaporate a few hundred pounds of acid per day, and were costly to replace, the logical thought presented itself of spraying the acid in a heated chamber. This new type of acid evaporator was simpler and more economical to operate and had sufficient capacity to fume several thousand pounds per day. The attached fig. 1 presents the outstanding features of the new furnace. A is the vaporizing drum, which is heated by the combustion gases of coal stoker fire box, B. Acid is sprayed into the hot gas stream within the drum through a number of air-acid atomizers, C, and the acid vapor, together with combustion gases, is introduced into the smoke stream through two cast iron pipes extending to different points in the flue. The operation of this type of furnace requires control of temperature inside the drum to prevent overheating, which dissociates the acid vapor to SO, and eventually to SO2, with resulting loss of acid. The dissociation to SO, at elevated temperatures is comparatively high, and in some cases accounts for as much as 60 pct of the acid sprayed into the drum. To reduce the loss of acid, through dissociation, some Cottrell plants, where the rate of acid sprayed reaches at times 9 ppm, use two evaporating units. Fuming the acid into the smoke stream is only

Jan 1, 1951

By Aladar B. Dory

INTRODUCTION Anyone familiar with the problems of hardrock mining will agree that the majority of the serious dangers present in mining are quite visible and obvious to any person reasonably familiar with the profession. Having unsecured, unscaled back over ones head, gives one a very good chance of ending up under a caved in mass of rock. Staying too close to a blast gives one almost a certainty of being hit by a flying rock. Too little oxygen in the air will very quickly lead to loss of consciousness and death. One walks only so much over deep, unsecured openings before he falls into them. It is because of this clear visibility of the conventional health and safety hazards that mining regulations in almost all jurisdictions world-wide are a more or less comprehensive collection of "shalls" and "must nots" of good common sense. When basic rules of common sense safe working practices are at stake, there is little room for dialogue and compromise. The mine inspector is then observing, during his inspection, how well the mine follows these common sense rules. RADIATION AS A HIDDEN DANGER Radiation in mines is a risk, the impact of which does not demonstrate itself immediately. It is first of all a potential risk. Two individuals exposed to identical radiation will almost certainly be effected differently, if at all. This is certainly true of exposures and doses one might encounter in the mines today. We hear very often the phrase: "there is very little known about the effects of radiation". It is one of the most misused and misunderstood half-true statements. I would doubt that there is any other carcinogen whose effects have been studied as extensively as the health effects of radiation. Where the statement is correct is regarding the knowledge of the quantitative assessment of the risk of low level radiation exposures. The reason for this uncertainty is that the magnitude of their health effect is very close to the health effects of natural radiation, cosmic radiation and the effects of other carcinogens such as industrial pollution, hydrocarbons from cars and other chemicals we have grown accustomed to using. As far as lung cancer is concerned, the effects of wide use of tobacco probably outperforms any other single substance. All this having been said, the bottom line is still unchanged. Radiation exposure, in most cases mainly radon daughter exposure, was and still is one of the health hazards of uranium mining and as such has to be controlled to the best of our ability. Various jurisdictions have adopted different approaches to the control of radiation exposures of uranium minemill workers. The following sections of this presentation will attempt to explain the regulatory approach taken in Canada. THE CANADIAN REGULATORY PHILOSOPHY As indicated earlier, the health effects of low level radiation are quantitatively not yet defined and no proven threshold of radiation exposure exists. The Atomic Energy Control Board&apos;s (the Board&apos;s) regulatory system is based on the basic assumption that there is no absolutely safe limit of radiation exposure below which there are no health effects. Theoretically we should therefore strive to reach zero exposure. It is obvious that this objective cannot be reached in real life. The objective of the regulatory process therefore has to be to achieve radiation exposures of the workers that are as low as reasonably achievable, social and economic factors taken into account. This, of course, is the internationally acclaimed ALARA principle put forward by the International Commission on Radiological Protection (ICRP). To avoid any misunderstanding it is worth emphasizing that the ALARA principle is applied to achieve exposures below the regulatory limits which must not be exceeded in normal operation of any nuclear facility including uranium mines and mills. The present regulatory limits for radiation exposures of atomic radiation workers are based on the recommendations of the ICRP and they are almost universally accepted. They should ensure that the risk from radiation exposure is not greater than the risk associated with working in a comparatively safe industry. Basically, there could be two extreme approaches to the regulation of uranium mining and milling. One extreme approach is to develop very extensive and detailed regulations and requirements covering all aspects of radiation protection. This is a somewhat autocratic approach to the regulatory process. This approach has two very serious shortcomings. If detailed requirements are set in regulations, due to the great variations of actual conditions at various mine-mill facilities, they have to be set as a compromise between the desirable requirements and those which could be met by practically all facilities. This approach takes away from the management of the facilities the initiative to strive for improved conditions. Requirements are spelled out in clear, understandable targets and the only worry of the management is to comply with these targets. One of the basic duties of management is to manage the operations in the most effective way with the maximum health and safety of the workers in mind.

Jan 1, 1981