Search Documents

By P. Somasundaran, D. W. Fuerstenau

This article presents the results of an analysis of grinding of 1:1 mixtures of 4x8-mesh quartz and limestone in laboratory ball and rod mills. From these experiments, the amount of energy consumed in grinding each component of the mixture was determined. In the ball mill, limestone consumes slightly more energy than quartz in the initial stages of grinding, but as the grinding proceeds quartz begins to consume a slightly greater fraction of the energy. The postulate that the fraction of energy consumed by any material in a ball mill is proportional to its volume fraction has been shown to be approximately true. On the other hand, in a rod mill energy consumption by quartz and limestone is equal only in the initial stages of grinding, with energy consumption by quartz greatly exceeding that by limestone as rod milling proceeds. Preferential energy consumption in tumbling mills has been discussed in a number of papers. For example, from an analysis of the kinetics of comminu- tion in ball mills, Arbiter and Bhrany' obtained rate equations based on the postulate that only coarse particles are ground, indicating that fine particles are protected from the action of the balls by coarser particles. In rod mills this protective action is much more pronounced than in ball mills. Recent studies on the rod milling of limestone-quartz mixtures have shown that this protective action causes the mineral with the lowest grindability (quartz in this instance) to consume the greater fraction of energy. 2 However, in the ball milling of mixtures it has been postulated that the consumption of energy by one of the constituents is determined by its volume fraction within the mill. 3 Possibly this hypothesis may be a simplification of the ball milling of mixtures since the protective action suggested by Arbiter and Bhrany may result in some preferential energy consumption by one of the constituents of a mixture. The objective of this study has been to determine preferential energy consumption in laboratory ball and rod mills by analyzing size distributions of each material in the products obtained from grinding 1:l mixtures of quartz and limestone. MATERIALS AND METHOD The procedure for carrying out the tests on wet grinding quartz-limestone mixtures has already been described.'r3 In the present investigation, 1:l mix-

Jan 1, 1963

By Thomas M. Barczak

The era of the shield-supported longwall face began in the United States a little over 25 years ago. The most fundamental development of the shield in the past 25 years has been an increase in size and capacity and a progression toward two-legged designs in favor over four-legged shields. Is the “bigger-the-better” design philosophy justified? The advantage of the two-legged design is largely attributed to its active horizontal force caused by the forward orientation of the leg cylinders. How do we know this? You might be surprised to learn that shield loading data may suggest otherwise. In addition, what about setting pressures? Should they be set as high as possible? Some people have advocated that setting pressures should equal the yield pressure. Does this make any sense? Is yielding a good or bad thing? These are some questions still worthy of investigation that are addressed in this paper. The first major change in gateroad support was the use of cable bolts to replace conventional wood cribbing particularly in Western mines. Standing support changed very little until the mid-1990s, when Strata Products1 introduced the Hercules and Propsetter supports in Eastern mines and Burrell Mining Products introduced the Can support in Western mines. Since then, there has been a dramatic increase in new roof support technologies for longwall gateroad support. Historically, standing support, including shield supports, have been designed based on a simplistic model requiring the support to have sufficient capacity to support the gravity loading of a perceived detached rock mass. This approach ignores the stiffness of the support and the resulting ground deformation that is known to be critical to roof stability. Recently, a new approach based on the ground reaction concept has been promoted by the National Institute for Occupational Safety and Health (NIOSH) to account for the interaction of the support with the rock mass behavior in achieving ground stability. In addition to gateroad support design, this concept has ramifications for shield capacities, setting pressures, and recovery room support design that challenge current practices in these areas. Although longwall mining has matured into the safest and most productive underground mining method used today, traditional concepts of roof support design and practice should still be evaluated and challenged. Ultimately, advancements in the science of ground control are made by recognizing deficiencies in current theories and thinking outside the box.

By Jérôme Girard, Shiu Kam, Sue Longo

"Historically, tailings management in a northern climate has involved the deposition of tailings into the Tailings Management Area (TMA) as conventional slurry. This type of system is obviously viable but presents challenges that are difficult to manage as circumstances change over the life of mine. These issues can include:? Increases in mill throughput, and? Life of the TMA for the life of mine.This case study will present how thickened, high-density, non-segregating tailings are used as an alternative to conventional slurry disposal to achieve greater storage capacity and enhanced environmental performance for a gold tailings facility in a northern climate. This paper will focus on how this change in deposition strategy accommodated the increased throughput of the mill and the increased life of mine. Specifically, the paper will discuss the addition and integration of new processes to the exiting mill system to accommodate the tailings management issues. It will highlight the challenges from laboratory testing to pre-feasibility to the engineered design of the thickened tailings system implemented at Musselwhite Mine.BACKGROUNDGoldcorp Canada’s Musselwhite Mine, in Northwestern Ontario (120 km north of Pickle Lake), is an underground gold mining operation that began production in 1997. The current mill throughput is about 4000 tpd (1.4 Mtpa). Gold extraction is based on a CIP process with cyanide destruction.The existing tailings management area (TMA) is located about 2 km southwest of the mill complex. The TMA is contained by high ground to the north and to the west and flanged by a moraine ridge along the south side. A number of dams have been constructed along this ridge. The dams were first built in 1996 and subsequently raised in 2000 and again in 2005. Excess tailings pond water is discharged to a polishing pond then released to the environment via a wetland.The TMA was originally designed to contain 13.7 Mt of tailings under a ponded surface for closure. The water cover was to inhibit adverse long-term environmental effects because the tailings contain pyrrhotite and other sulphide-bearing minerals that are potentially acid generating. The design was modified during Stage 3 Construction in 2005 to expand the storage capacity to 17.3 Mt (Golder, 2005). As of end of 2009, over 16 Mt of tailings had been deposited in the TMA. The slurried tailings were historically deposited at 50% solids."

Jan 1, 2011

By Baoquan An

The energy required to fragment a rock is proportional to the increase in the fracture surface area and hence is proportional to the rock?s surface area. In comminution, the slope of the linear relationship between size reduction ratio and specific energy is usually defined as grindability. Another definition of grindability based on newly created specific surface area and specific energy is introduced. The two definitions for grindability reflect the grindability properties of rocks and can be used to measure of the efficiency of mills or crushers. The feed particle size distribution influences the energy needed to comminute the rock. For a given ratio of average feed size to product size, the smaller feed size requires more energy.

May 1, 2004

By James D. Nickerson

Thee Brightwater water treatment facilities, currently scheduled to be completed in 2010 and begin operating in 2011, will serve the northern part of King County, WA. The facility includes a new treatment plant and extensive conveyance system consisting of more than 21 km (13 miles) of influent and effluent conveyance tunnels and five portals. The majority of the conveyance system tunnels are mined by tunnel boring machines (TBM) and microtunnel boring machines (MTBM). Tunnel and portal construction is divided into three main contracts. The Ballinger Way receiving portal, constructed under the central tunnel contract, is the terminus for the BT-3 leg of the central contract and the BT-4 terminus of the west contract. The shaft has a 7.3-m (24-ft) finished inside diameter (ID) to accommodate removal of the central project TBM and the west contract TBM and extends to a depth of 66 m (216 ft) below grade through a complex and challenging soil profile characterized by multiple ground water tables. Vinci Grand Projects/Parsons/Frontier-Kemper Constructors (VCFK) was the joint venture team for the Ballinger Way portal. Frontier-Kemper was responsible for the shaft sinking. VCFK evaluated either slurry wall techniques or ground freezing for earth support during excavation of the portal and liner installation. The selection process took into account the vertical depth, cost, past practice, geology, water tables and a 9-m- (30-ft-) thick zone of contaminated soil at the top of the shaft. Ground freezing, which would also act as a hydraulic barrier, was deemed the better option for this project. The contract was awarded to New Jersey-based geotechnical contractor Moretrench. Typically, frozen walls, by design, key into an underlying cutoff. At the Ballinger Way portal, however, the [ ] freeze bottomed out in a substantially thick stratum consisting of predominantly clay, but with some inclusions of permeable soil that indicated significant ground water pressures. To ensure stable conditions at the bottom of the shaft, Moretrench, which also specializes in dewatering, designed and installed a perimeter array of deep wells to temporarily depressurize the permeable soils. The movement of ground water generated by pressure

Jan 1, 2009

By S. H. Katz

"At the annual meeting of the International Association of Fire Engineers at Kansas City, Mo., in June, 1919, the Bureau of Mines offered to cooperate with municipal fire departments for the purpose of obtaining information regarding composition of the air in burning buildings. (Katz, S. H., Statement made at Convention of International Assoc. of Fire Engineers: Fire and Water Eng., vol. 66, July 16, 1919, pp. 129-30, also Fireman's Herald, vol. 78, July 12, 1919, pp. 48.) The gas-mask laboratory of the Bureau's experiment station at Pittsburgh sent vacuum bottles for taking samples to fire chiefs in twenty-five cities, who had signified their intention to aid in this work. It was hoped to obtain one hundred or more samples, because statistics in quantity are necessary for drawing reliable conclusions concern-in the likelihoods of encountering irrespirable as in burning buildings. This far, eleven samples have been received by the laboratory. Results of analyses and pertinent data regarding these samples are given in the accompanyingEight of the eleven samples contained, no carbon monoxide; seven of the eight were taken above the ground level and one below ground. One sample taken above ground contained 0.03 per cent carbon monoxide, and this quantity is not dangerous to breathe for a limited time. Of the three samples taken below ground, one contained no carbon monoxide; the other two contained 0.10 and, 0.35 per cent, respectively. This indicates that in a burning building, the hazard from carbon monoxide is much greater below the ground level. The cause undoubtedly lies in the poorer ventilation that is usually found in cellars or basements. Abundance of ventilation provides the necessary air for complete combustion, or if incomplete combustion does occur, the carbon monoxide generated by the fire is greatly diluted with fresh air and swept away by the current. It may be noted that the highest proportion of carbon monoxide was found in a cellar where a fire as burning among piled cases. The difficult access of air to the burning parts facilitated the formation and lib ration of carbon monoxide. The smoke here was very light and no difficulty was experienced in breathing. Some of the air containing dense smoke showed little change from normal air by analysis. Thus the density of smoke or the irritation and difficulty it causes in breathing are no indication of the presence of poisonous gas. None of the samples shows important oxygen."

Nov 1, 1920

By Alastair Grogan

In October of 2002, the author was retained by Public Works and Government Services Canada, on behalf of the Department of National Defense, to design, construct and test the effectiveness of a bubble curtain system at an underwater demolition range, near Victoria, BC. The purpose of the bubble curtain was to reduce the potential for fish mortalities at the site. This paper presents some of the elements of the design, along with the results of overpressure monitoring during blasting and compares the results of fish mortality and animal behaviors with and without the bubble curtain in operation.

Jan 1, 2005

By N. M. Skarbøvig

Safety nets have been used selectively in several South African mines over the years. Recently, however, an increased emphasis on mine safety has seen a resurgence of their utilization underground. This renewed interest has, however, also raised questions about the design and performance of these nets and their practical applications underground; these are the subjects of this paper. Generally, mine safety nets are used to protect personnel from falling rock in temporarily supported development ends, tunnels, ASGs and stopes. These diverse applications each require specific net designs, resulting in a wide product range. As a rule, however, mine safety nets need to carry specified static or dynamic loads with the least possible deformation. Because of the application-specific characteristics of mine safety nets, there is no South African national standard for them. Consequently, neither the SABS, nor the CSIR, performs standard tests on these components. In 2007, therefore, SIMRAC made available their Savuka facility for the testing of mine safety nets; the Savuka drop test became the accepted industry test standard for mine safety nets. From late 2007 to mid 2009, Norsenet, in collaboration with SRK Consulting, the designated operators of the site, conducted approximately 200 tests, which formed the core of Norsenet's database of safety net performance under various loading conditions. The test specimens comprised a wide range of net materials and config¬urations; the materials included: high density polyethylene, polypropylene, polyester, nylon, as well as a blend of polyester and polypropylene used primarily for reinforcement ropes. Net configurations encompassed the following parameters: net type, i.e., knotted nets or sewn nets; net strand type, i.e., braids, twines, or webbing; strand size; mesh size; mesh orientation, i.e., parallel or oblique to the net edge; net edge reinforcement, i.e., knotted, tied, with or without reinforcement ropes; internal reinforcement ropes, i.e., diagonal ropes and/or orthogonal ropes; as well as net attachment options, including loops and chainlinks. From the number of variables listed above, and the limited number of tests available, it can be appreciated that the findings of the test series required careful scrutiny. Nevertheless, principal performance factors clearly emerged: besides the obvious net mesh variables, such as net strand material, strand diameter and mesh size, it was the configuration of the reinforcement ropes and their attachment to the net panel which had the strongest influence on net performance. All the structural elements of a safety net need to work together to transfer the imposed load from the net panel, through the reinforcement ropes, to the net attachment points. In this load transference chain, the weakest link limits the strength of the structure. From this development work, Norsenet has gained valuable expertise to custom-design and manufacture nets to clients' specifications. The examples presented in the paper will also show the experience acquired in the installation of nets underground. Future work will include the establishment of a state-of-the-art net testing facility at Norsenet's factory in Vredenburg.. Keywords Mine safety net, safety net, protective device, mine safety, safety net design, safety net performance, safety net applications.

Jan 1, 2011

By F. -J. Wodopia

Sustainability of energy supplies has been an issue of general interest for a long time. Whilst in the past the continuous productive use of natural commodities was in the fore, its compatibility with the environment and climate and ultimately also social aspects have recently gained more and more attention. Today many discussions focus on the climate compatibility of individual options to act. Fossil fuels, in particular coal and lignite, are generally challenged because coal use makes a significant contribution to global man-made CO2 emissions. The drastic increase in the consumption of fossil fuels is not explained by the fact that coal utilisation technologies are not or cannot be sustainable but rather by the growing world population and the endeavours numerous national economies trying to catch up economically. Coal is the world’s No. 1 fuel and, in the view of many experts, will keep this position also in the foreseeable future. The access to its deposits plays a significant role for the economies of the emerging countries. Consequently the aspects of energy security and economic efficiency need to be considered at eye level with climate compatibility. In Germany, however, the public acceptance of fossil fuels and of large industrial projects is in the focus of attention.

Jan 1, 2011

By John Henderson

Densities of melts 0f the lime-iron oxide-silica system in contact with solid iron have been measured by the maximum bubble pressure method in the temperature range 1250° to 1440°C and the composition range 0 to 40 mol pct lime, 15 to 100 mol pct iron oxide, and 0 to 55 mol pct silica. Densities range from 4.65 g cm 3 for wustite at 1440°C to 2.75 g cm-&apos; at 1350°C for a melt containing 30 mol pct lime, 20 mol pct iron oxide, and 50 mol pct silica. The results are interpreted in terms of a postulate that the melts can be regarded as a random array of oxygen ions in which regions of local order exist to satisfy the coordination requirements 0.f the cations. An understanding of the nature of metallurgical slags is basic to the development of a sound theoretical description of heavy metallurgical extractive and refining processes. Because these liquids are complex, direct measurements of their properties has not thrown much light on their structure. This has led to the approach of measuring the properties of simpler liquids, and building up their complexity until slag compositions are reached. In this way the density of liquid iron silicates was measured in a previous study1 and the present work represents a further stage in this synthesis. EXPERIMENTAL The technique used in the measurement of density was the maximum bubble pressure method. Details of the apparatus and procedure were similar to those previously reported,&apos; with the exception that a constant voltage transformer was used to supply the power input to the furnace and six silicon carbide resistance elements were used in place of the molybdenum winding. With these modifications melt temperature could be maintained within 1 centigrade degree during the course of a run. The silica used to prepare the melts was washed natural quartz ignited at 1000°C; wustite was prepared by air-melting A.R. grade ferric oxide in an iron crucible and lime was prepared by air ignition, at 1000°C, of weighed quantities of A.R. grade calcium carbonate, previously air-dried at 110°C. The finely ground constituents were intimately mixed in a glass ball mill prior to melting. Temperatures quoted are accurate to * 5°C and the standard deviation of the density values, calculated by the method of least squares, ranged from 0.5 to 1.8 pet. However, replicate determinations of density on different melts of the same nominal composition at the same nominal temperature did not vary by more than 1 pct, Table I, and this figure has been taken as an estimate of the accuracy of the density results. The density of carbon tetra-chloride was also measured as a check on the absolute performance of the experimental method. At 20°C a value of 1.593 * 0.002 g cm"3 was obtained; this compares with the literature value2 of 1.595 g cm"3. Results of experiments designed to measure the dependence of the density of lime-iron oxide-silica melts, in contact with solid iron, on composition and temperature are shown in Table I. Because iron sometimes precipitated in the sample during quenching, the Fe203 chemical analyses were only poorly reproducible and should be taken as a guide rather than as absolute values. Fig. 1 shows the data from various sources for the density of liquid iron silicates and Fig. 2 shows isodensity contours at 1410°C for lime-iron oxide-silica melts, calculated by graphical interpolation of smoothed curves drawn through the experimental results, together with the 1400°C results of Adachi and ogino3 and Pope1 and Esin.4 Fig. 3 shows the isothermal variation with composition of the volume of melt per gram ion of oxygen at 1410°C and Fig. 4 shows regions in which the temperature coefficient of this volume is negative, positive, or negligible (<0.005 cm3 deg-I). DISCUSSION a) Disparity Between Reported Density Results. Consider the system iron oxide-silica, the results for which are summarized in Fig. 1. Although there is some difference in the temperatures at which the various densities apply, this difference is not sufficiently large to account for the observed discrepancies. The reliability of the present results for the low-silica region has been confirmed by measurement of the density of liquid wustite by three different techniques. At 1410°C the density measured by a balanced-column method was 4.55 g cm"3, by a combination balanced-column and gas-densitometer method 4.59 g emd3, and by a pycnometer method 4.53 g cm"3. Schenck, Frohberg, and Hoffermann&apos; have also reported a value of 4.55 g cm"3 for the density of liquid wustite at 1400°C. It must be concluded, therefore, that neither Pope1

Jan 1, 1964

By Rollyn P. Jacobson

THE town of Pacific, in Jefferson County, Mo., is 127 miles west of St. Louis. Since the area lies entirely on the flood plain of a cutoff meander of the Meramac River, it was considered a likely environment for accumulation of commercial quantities of sand and gravel. Excellent transportation facilities are afforded by two major railways to St. Louis, and ample water supply for washing and separation is assured by the proximity of the river. As a large washing and separation plant was planned, the property was evaluated in detail to justify the high initial expenditure. An intensive testing program using both geophysical and drilling methods was designed and carried out. The prospect was surveyed topographically and a 200-ft grid staked on which electrical resistivity depth profiles were observed at 130 points. The Wenner 4-electrode configuration and earth resistivity apparatus" were used. In all but a few cases, the electrode spacing, A, was increased in increments of 11/2 ft to a spread of 30 ft and in increments of 3 ft thereafter. Initial drilling was done with a rig designated as the California Earth Boring Machine, which uses a bucket-shaped bit and produces a hole 3 ft in diam. Because of excessive water conditions and lack of consolidation in the gravel there was considerable loss of hole with this type of equipment. A standard churn drill was employed, therefore, to penetrate to bedrock. Eighteen bucket-drill holes and eight churn-drill holes were drilled at widely scattered locations on the grill. The depth to bedrock and the configuration will not be discussed, as this parameter is not the primary concern. Thickness of overburden overlying the gravel beds or lenses became the important economic criterion of the prospect.** The wide variety and gradational character of the geologic conditions prevailing in this area are illustrated by sample sections on Fig. 2. Depth profiles at stations E-3 and J-7 are very similar in shape and numerical range, but as shown by drilling, they are measures of very different geologic sequences. At 5-7 the gravel is overlain by 15 ft of overburden, but at E-3 bedrock is overlain by about 5 ft of soil and mantle. Stations L-8 and H-18 are representative of areas where gravel lies within 10 ft of surface. In most profiles of this type it was very difficult to locate the resistivity breaks denoting the overburden-gravel interface. In a number of cases, as shown by stations M-4 and H-18, the anomaly produced by the water table or the moisture line often obscured the anomaly due to gravel or was mistaken for it. In any case, the precise determination of depth to gravel was prevented by the gradual transition from sand to sandy gravel to gravel. In spite of these difficulties, errors involved in the interpretation were not greatly out of order. However, results indicated that the prospect was very nearly marginal from an economic point of view, and to justify expenditures for plant facilities a more precise evaluation was undertaken. The most favorable sections of the property were tested with hand augers. The original grid was followed. In all, 46 hand auger holes were drilled to gravel or refusal and the results made available to the writer for further analysis and interpretation. When data for this survey was studied, it immediately became apparent that a very definite correlation existed between the numerical value of the apparent resistivity at some constant depth and the thickness of the overburden. Such a correlation is seldom regarded in interpretation in more than a very qualitative way, except in the various theoretical methods developed by Hummel, Tagg (Ref. 1, pp. 136-139), Roman (Ref. 2, pp. 6-12), Rosenzweig (Ref. 3, pp. 408-417), and Wilcox (Ref. 4, pp. 36-46). Various statistical procedures were used to place this relationship on a quantitative basis. The large amount of drilling information available made such an approach feasible. The thickness of overburden was plotted against the apparent resistivity at a constant depth less than the depth of bedrock for the 65 stations where drilling information was available. A curve of best fit was drawn through these points and the equation of the curve determined. For this relationship the curve was found to be of the form p = b D where p is the apparent resistivity, D the thickness of overburden, and b a constant. The equation is of the power type and plots as a straight line on log-log paper. The statistical validity of this equation was analyzed by computation of a parameter called Pearson&apos;s correlation coefficient for several different depths of measurements, see Ref. 5, pp. 196-241. In all but those measurements taken at relatively shallow depths, the correlation as given by this general equation was found to have a high order of validity on the basis of statistical theory.

Jan 1, 1956

By E. J. Rapperport, J. P. Pemsler

The thermodynamic activities of zinc in six solid solution Cu-Zn alloys ranging from 5 to 35 at. pct Zn were determined experimentally in the temperature range 400° to 600°C. This low temperature investigation was canducted in order to evaluate techniques developed to utilize the inherently high sensitivity of atomic absorption flocesses in the measurement of thermodynamic activities. Analytical expressions ,for the activity and actizlity coeflcient are given for the six alloys in the temperature ranges investigated. RELATIVELY few experimental methods are available for investigation of thermodynamic activities of alloys, especially in the solid state. The techniques most frequently used have been the electrochemical potential and the effusion methods, both of which have severe limitations in many instances. It is therefore desirable to expand the ability to perform such measurements in order to obtain new information as well as to provide an additional independent verification capability. In this work, we present a significant improvement in the spectrophotometric method for sensing small vapor pressures in static absorption cells. Similar techniques have been used previously;1"5 however, applications had been limited to relatively high pressures, often greater than 1 torr. Prior investigators have, for the most part, used broad spectral sources such as xenon or mercury lamps, and high intensity arcs. Hollow cathode sources were first suggested in 1956 6 and were used soon afterwards.4&apos;5 These sources offer significant improvements in sensitivity and freedom from interfering spectral lines.&apos;-&apos; EXPERIMENTAL High purity zinc was obtained from Cominco Products, Inc., and copper from American Smelting and Refining Co. Both elements were of 99.999 pct purity. Copper-zinc alloys were vacuum melted in a high fired carbon crucible with each alloy pulled from the melt as a 4 -in. diam bar. The bars were swaged to -1/4 in. rods and vacuum annealed for 160 hr at 800° + 1°C. Samples for gross chemical analysis were taken at intervals along the length of the rods to ascertain the axial zinc gradient. Electron microprobe analysis of homogenized specimens indicated that the alloys had uniform compositions over their cross sections on a macro (200 p) and micro (1 u) scale to better than *1 pct (20) of the gross composition. This tolerance was determined by counting statistics, rather than assured composition fluctuations. All SiO 2 windows were high-ultraviolet-transmission grade to minimize intensity losses. Silica absorption cells were scrupulously cleaned consecutively in organic solvents, dilute HF, and distilled water before use. The empty cells were then flamed while under a dynamic vacuum, cooled, and removed to an argon-filled glove bag. Alloy pieces were cut and filed in the glove bag to produce fresh surfaces, and then loaded into the cells. The loaded cells were temporarily sealed, removed from the glove bag, reevacuated to 10-5 torr or better, and permanently sealed. The instrument used is schematically shown in Fig. 1. The spectral emission from a commercially made hollow cathode lamp (A) of a selected element is focused through an absorption cell (B) inside a well-controlled furnace (C). The intensity of the transmitted beam is measured using the spectrometer* (D) &apos;Techtron model AA4 atomic absorption spectrometer. which contains a grating (E) that disperses the light prior to impingement on the photomultiplier (F). The monochromator grating is adjusted so that only the wavelength of interest is measured. The power supply delivered an interrupted voltage to the lamp, causing a chopped radiation output to be transmitted. The detector read only the intermittent component of radiation incident upon it, so that all continuous noise signals (furnace radiation, and so forth) were eliminated. Three recording thermocouples contained in the muffle furnace were positioned along the length of the absorption cell: one at each end and one at the center. An effort was made to keep the ends of the cell several degrees hotter than the center to avoid window condensate. Appropriate thermal corrections were then necessary to relate cell pressure to radiation attenuation. Water-cooled heat shields, as shown in Fig. 1, were found to aid signal stability by protecting the hollow cathode and the photomultiplier from furnace radiation. The furnace had a 2-in. diam muffle, Kan-thal wound, with SiO 2 windows at its ends to minimize convective effects. The hollow cathode radiation was masked and focused to form a conic beam that was a maximum of { in. diam within the furnace. Thus, the 1.5 in. diam absorption cell easily contained the entire beam. The furnace was mounted on ball-bearing slides with positive positioning detents. This arrangement allowed the removal of the entire furnace assembly from the radiation path, position [I], Fig. 1, so that frequent sampling of the unattenuated beam intensity could be obtained. In all cases the beam intensity was kept constant to 0.1 pct as judged by readings taken immediately before and immediately after data collection. Only data for absorptions of less than 80 pct were utilized, as systematic deviations from linearity were found for greater absorptions.

Jan 1, 1970

By T. Y. Yan

SUMMARY Laboratory high pressure column tests have shown that the presence of 1-20 ppm of aluminum ion effectively prevents permeability loss during uranium leaching with leachates containing sodium carbonate. If added after permeability loss has occurred, aluminum ion can restore the permeability to nearly its original value. No deleterious effect was observed on uranium leaching performance and the technique should be quite compatible with all field operations. INTRODUCTION The recovery of uranium values from underground deposits by in situ leaching or solution mining has become economically viable and competitive with conventional open pit or underground mining/milling systems (Merrit, 1971). In situ leaching processes are particularly suitable for small, low-grade deposits located in deep formations and dispersed in many thin layers. Many such ore bodies occur along a broad band of the Gulf Coastal Plain (Eargle et. al., 1971). The advantages of the in situ leaching processes have been reviewed (Anderson and Ritchi, 1968). In the in situ leaching process, a lixiviant containing the leaching chemicals is injected into the subterranean deposit and solubilizes uranium as it traverses the ore body. The pregnant lixiviant or leachate is produced from the production well and is then treated to recover the uranium. The resulting barren solution is made up with the leaching chemical to form lixiviant for re-injection. Upon completion of the leaching operation, the formation is contaminated with leaching chemicals and other species made soluble in the leaching operation and has to be treated to reduce the concentration of these contaminants in the ground water to levels acceptable to the regulatory agencies (Witlington and Taylor, 1978). Restoration is accomplished by injecting a restoration fluid, which could be the fresh water or water containing chemicals, into the formation. As it traverses the leached formation, the restoration fluid picks up the contaminants and is then produced at the production well. This produced water is either disposed or purified for recycle. In both phases of operation, formation permeability or well injectivity is one of the most important parameters which determines the viability of the in situ leaching process. Low formation permeability limits production rates, leading to uneconomical operations. The formation is said to be sensitive if there is a sharp loss of permeability on contact with water and other fluids. Many uranium bearing formations, for example, the Catahoula formation of the Texas Coastal Plain, contain significant amounts of clay minerals which are water sensitive. Serious permeability losses can occur when the pH and chemical composition of the lixiviant is significantly different from that of the formation water. Jones has investigated the influence of chemical composition of water on clay blocking of permeability (Jones, 1964) and Mungan studied permeability reduction through changes in pH and salinity of the water (Mungan, 1965). Various mechanisms of permeability damage have been proposed and reviewed (Jones, 1964; Mungan, 1965; Gray and Rex, 1966; and Veley, 1969). When large amounts of swelling clays are present, a significant fraction of the flow channels in the formation can be reduced due to swelling. However, in most cases, swelling need not be the main cause of permeability losses. Particle dispersion and migration or clay sliming can be more important causes for formation damage. Clay particles entrained in the moving fluids are carried downstream until they lodge in pore constrictions. As a result, microscopic filter cakes are formed by these obstructions, plugging the pores, effectively restricting fluid flow and reducing the formation permeability. Moore found that as little as 1-4 percent clays present in a fine grained sandstone could completely plug the formation if they are contacted by incompatible injected fluids (Moore, 1960). It has been found that injection of NaHC03/Na2CO3 lixiviant into formations with significant clay content often leads to loss of formation permeability and well injectivity. To alleviate this problem a change of the lixiviant composition to KHC03/K2CO3 has been proposed. At present, however, many in situ leaching operations employ NH4HC03/(NH4)2C03 mixtures as a source of carbonates. This approach has been successfully used in South Texas by Mobil, Intercontinental Energy, Wyoming Minerals and U.S. Steel, etc. The use of ammonium carbonates solutions, however, contaminates the formation and requires a time-consuming restoration operation. The other approach to reduce the permeability loss is to pretreat the sensitive formation with chemicals which prevent clay dispersion and migration. Such chemicals include hydroxy-aluminum (Reed, 1972 and Coppel et. al., 1973), hydrolyzable zirconium salts (Peters and Stout, 1977), hydrolyzable metal ions in general (Veley, 1969) and polyelectrolyte polymers (Anonymous). Still another approach, is to minimize the "shock" caused by sudden injection by gradually changing the chemical composition of the injected fluids from that of the formation water. THE APPROACH Since permeability loss can be an important factor limiting the efficiency and economic viability of the in situ leaching process, a study was initiated on

Jan 1, 1982

By J. Brett, L. Seigle, L. Castleman, T. Montelbano

Impurity-induced low-temperature recrystallization of cold-worked tungsten was inuestigated with emphasis on the influence of nickel on the reaction. Palladium, nickel, aluminum, manganese, platinum, and iron greatly lower the recrystallization temperature of doped tungsten, which zs normally very high, but the recrystallization temperature of electron-be am zone-refined tungsten wzre is slightly raised by conlacl with nickel. Recrystallization can be induced at low temperature by the presence of solid nickel on the surface of doped tungsten wire, but apparently not by exposure to nickel vapors alone. Approximately 200 ppm of Ni dijjused into 10-mil wires at 1200 from a deposit of nickel on the surface produced total recrystallization, whereas more than 600 ppm of Ni could be absorbed frotn a vapor source without altering the fibrous structure of cold-worked tungsten. Once initiated, nickel-induced recrystallization required a continued source of&apos; nickel for propagation of the recrystallization front. The solubility of nickel in fibrous 10-mil W wire was approximalely 500 ppm at 1150&apos; C, and the activation energy for penetration of the recrystallization front was 52 kcal per mole. In many applications the usefulness of tungsten depends on critical control of its structure. Cold-worked tungsten with the fibrous structure developed by suitable thermo-mechanical treatment has a low, but technologically significant, ductility. It has long been known1 that traces of nickel, and perhaps other metals, are profoundly deleterious in doped tungsten, because they induce recrystallization at low temperature, which produces a brittle, equiaxed grain structure. This effect appears to be an exception to the general observation that recrystallization is impeded and the recrystallization temperature raised by the presence of impurities.2"9 Previous studies7-&apos;&apos; of the annealing and recrystallization of tungsten wire have divided the phenomenon into prior recovery stages, primary recrystallization and secondary recrystallization. The present investigation is concerned principally with primary recrystallization which is defined here as the replacement of the fibrous structure of deformed tungsten by equiaxed grains. The objective of this study was to explore the nickel-induced recrystallization reaction in tungsten and attempt to elucidate its mechanism. As well, an effort to define which other elements give rise to low-temperature-induced recrystallization was carried out. EXPERIMENTAL PROCEDURE The procedure adopted for these experiments was essentially to bring nickel and other elements into diffusive contact with cold-worked tungsten wires. The process of recrystallization was followed as a function of time and temperature by light and electron microscopic observations. First the influence of nickel on the recrystallization temperature of arc-melted, zone-refined, and variously doped tungsten wire was determined by electroplating a deposit of nickel on the surface of the wire and annealing at a variety of temperatures for 3 hr. The chemical analyses of the tungsten wires used in this investigation are given in Table I. The surface of the tungsten wire was etched with Murakami&apos;s slution&apos; and approximately 0.005 in, of Ni was deposited from a Watts-type low pH bath14 for the conditions of these experiments. Variations in plating thickness from about 0.001 to 0.005 in. had no discernible influence on the resulting structures. The wires were then annealed in an atmosphere of dry hydrogen to establish the recrystallization temperature. Concurrently, un-plated specimens were annealed to establish the recrystallization characteristics of nickel-free wire. The criterion of recrystallization was that the fibrous structure be completely replaced by equiaxed grains after a 3-hr treatment of temperature. This provided more reproducible results than use of the first recrystallized grain or a fixed proportion of re-crystallized structure as the critical observation. The structures encountered in longitudinal and transverse sections were examined by both light and electron microscopy at magnifications up to X32,000 using parlo-dion-carbon replicas shadowed with platinum for the latter method. Second, the influence of a variety of metals on the recrystallization temperature of 0.010 in. D alumina-silica doped tungsten wire, AW136-64, was determined. The elements were applied by electroplating whenever possible. Alternatively, they were vapor-plated on the tungsten wire and a greater thickness built up by coating with a dispersion of metal powder in nitrocellulose lacquer. Elements not amenable to either of these procedures were merely slurry coated on the tungsten. The recrystallization temperature was determined as above. Third, the nickel-induced recrystallization process in doped wire was studied more closely by electroplating 8 mils of Ni on 65-mil alumina-silica doped tungsten wire, AW153-NS10, and exposing the wire to temperatures of llOO°, 1200°, or 1300°C for various times in a hydrogen atmosphere. A circular recrystallization front, &apos;Onsisting of equiaxed grains, developed at the periphery Of the coated tungsten wire, and the advance of this front into the fibrous interior was studied. These experiments employed relatively coarse 0.065 in. D wire because the 0.010 in. D wire recrystallized too quickly to permit observation of the pene-

Jan 1, 1969

By R. A. Davidson, L. Silverman

RESIDENTS of many urban centers are becoming increasingly aware of the obscuring effect of fume and smoke discharge from power, metallurgical, chemical, and other industries; and they, as well as the legislatures of these affected cities, are agitating for cleaner air. Management&apos;s most pressing problem is to find an economical way to reduce process effluents in response to the growing pressure from population and legislative demands. The removal must be done, if possible, without handicap to the current operation, since the costs of relocating are often excessive or prohibitive. In fume recovery or disposal, an important item to consider is whether or not the material being discharged has any value. If it has commercial value, the cost of its recovery may offset or aid amortization. For this reason, in making a study of the specific problem in hand, a major factor was the nature of the material emanating from the stack: in particular, its particle size, size range, and its chemical and physical composition, as well as its potential value and utility when recovered (in either a wet or dry state). Should the product have no commercial value, it must be disposed of at minimum cost in a way to prevent recontamination. Initial studies were therefore made to determine stack concentrations and volumes of material evolved from the operations. The next phase of the study concerned the physical and chemical nature of the collected fume. The third portion of this paper describes the wet and dry collector studies undertaken to recover the fume. Cleaning Requirements for Ferroalloy Furnace Operation The basic need for any effluent collection equipment is the highest possible efficiency and the lowest tolerable resistance when the power consumption involved is considered. Since the electric furnace effluent is largely composed of fume of small size (less than 0.5u), it has high light obscuring properties, and even low concentrations will cause some loss of visibility and be evident to nearby residents. The permissible limit for fly ash emission in many cities is based on a weight value (viz, approximately 0.4 grains per cu ft), but the smoke density values are dependent upon a shade of color. In the case of the Los Angeles County code, emission is restricted to pounds per pound of material processed per hour basis (but not exceeding 40 lb per hr for any one given plant operation). If an average particle size of the fume from ferro-silicon alloy electric furnaces is assumed to be 0.4u (as shown later, this is the approximate mean size) and an average loading of 1 grain per cu ft (stp), each cubic foot of stack gas will contain approximately 75x10 10 particles (based on assumed, and confirmed, spherical shape and a standard deviation of unity). When it is realized that the air in metropolitan areas, which are also general industrial areas, contains approximately 5x108 particles, the tremendous light scattering effect of this concentration becomes apparent. Consequently, nearly 100 pct collection would be necessary to equal the average concentration. Fortunately, however, discharge from a high point above ground (50 to 100 ft) will result in at least a thousandfold dilution, or the stack concentration reaching the ground in the foregoing case might result in a ground concentration of &apos; particles. If the concentration at the source could be reduced by a factor of 100 (99 pct efficiency of collection), then a concentration of 75x10" particles would be diluted to 7.5x10&apos; which would be very satisfactory. An efficiency of 90 pct (factor of 10 decontamination) at the source would result in a discharge of 75x109 articles which upon dilution yields 75x10 which is still 15 times the general air value. Another approach to this consideration is to use the value of concentration of 0.005 grains per cu ft for the value of a visible effluent as cited by Kayse.1 To attain this value with an average loading of 1 grain per cu ft would require an efficiency of 99.5 pct. Since the foregoing value is not based on any reported size of fume particles, it is felt that the numbers&apos; approach given previously is more reliable. These calculations serve to indicate the desirability of thorough cleaning, preferably at the source, and with efficiencies well above 90 pct, preferably above 95 pct (dilution 1:20). One of the most important items in any control program is to reduce the concentrations as close to their sources as possible. The use of better furnace design, deeper coverage over the electrodes, and the prevention of blows or breaks in the surface all help to reduce dissemination; consequently, all of these improvements should be made, if possible, to cut down the effluent load. In addition, in order to minimize the volume of contaminated air that has to be cleaned, the furnace should be enclosed as much as possible. Test Arrangements Before fundamental studies with collectors were made, a furnace stack selected for the test program was sampled to determine the gas temperatures and

Jan 1, 1956

By S. F. Reiter

The paper presents isothermal recrystallization curves for 0.08 and 0.15 pct C steel at subcritical temperatures following small amounts of plastic deformation. The effects of deformation, temperature, and aging on nucleation and growth rates ore described. The free energy of activation for grain boundary migration in steel is given. SEVERAL excellent reviews of the literature have appeared concerning the recrystallization of metals.&apos;-&apos; The present investigation follows the approach advanced by Mehl, Stanley, and Anderson,6-7 in which the rate of recrystallization was analyzed in terms of N, the rate of nucleation, and G, the rate of growth of recrystallization nuclei. Two lots of low carbon, capped steel of the analysis given in Table I were studied. Each lot consisted of a 150 lb coil which had been hot rolled to 0.083 in. and then cold rolled to 0.042 in. at the mill. Strips 0.930 in. wide were sheared perpendicular to the rolling direction. Both steels were normalized before studying their recrystallization characteristics. The strips were cleaned, painted with a magnesia-acetone paste, and made into packs of equal weight, wrapped in 0.002 in. copper foil. The packs were placed in a salt bath at 900°C for 30 min and air cooled. A relief anneal followed in a second salt bath for 15 min at 650°C. The relief anneal was found necessary from early tests in which a longer incubation period and slower rate of recrystallization were observed in relief-annealed lot A steel than in similar material which was strained and recrystallized directly after being normalized. This effect, which indicates the presence of transformation and/or cooling stresses in steel air cooled from above the A, temperature, has also been observed by Samuels8 and Masing.9 Figs. 1 and 2 show the microstructure of lot A and B materials and illustrate the rather uniform No. 8 ASTM grain size produced by this heat treatment. Winlock and Leiter10 observed that strip specimens which had their sharp edges removed elongated more uniformly than those which were not polished. Similarly, when the sheared edges were removed on a belt grinder, it was found in the present investigation that such samples recrystallized more uniformly than did unpolished strips. Therefore, all strips were carefully rounded prior to their extension. The approximate strain limits for the production of large recrystallized grains are from 6 to 12 pct extension." It was found that for the purpose of this investigation, 8 and 9 pct elongation were suitable deformations. The strain rate employed was 0.01 in. per in. per min and produced a yield point elongation of 4 pct. Winlock and Leiter found that mild steel of No. 8 ASTM grain size gave the same yield point elongation when extended at 0.012 in. per in. per min. All of the lot A and B strips extended in tension developed a straight, stretcher strain line at each grip when the upper yield point was reached. The lines were parallel and made an angle of 55" with the edge of the strip. They approached each other with increasing strain and met near the center of the sample at the end of the yield point elongation. Immediately thereafter, a small drop in load was observed and then the load increased in a regular manner with increasing extension. The grips were initially 8 in. apart. After extension, the 6 in. gage length was carefully cut into 1 in. samples. The remainder of the strip was discarded. After a flash pickle in hot 50-50 hydrochloric acid, six samples, each of which had been taken from a different strip, were placed in a basket and submerged in a lead pot for isothermal recrystallization. Although no recovery effect was observed, strain aging did occur after extension. Therefore, samples were always recrystallized within 24 hr after their cold deformation. After recrystallization, the samples were etched with a solution comprised of one part by volume of nitric acid with three parts of water. Bromide printing paper was exposed directly at low magnifications and later used with a mask to measure the desired quantities. First, the average diameter of the largest grain visible in each sample was determined using dividers. Next, the number of recrystallized grains per unit area was counted and recorded as n. Then, for each sample, the combined area of the recrystallized grains was measured by transcribing the grain outlines to standard graph paper. Many determinations of the area of the recrystallized grains were repeated five times and indicated a standard error that was not greater than 25 pct. The average area for six samples was divided by the area of the mask to yield the percentage recrystallized. Recrystallization of 0.08 Pct C Steel The progress of recrystallization at 670°C after 8 pct elongation of lot A steel is shown in Fig. 3, a through f. The shapes of the growing crystals are approximately equiaxed, as is assumed in the

Jan 1, 1953

By M. E. Wadsworth, K. C. Bowles, H. E. Flanders, R. M. Nadkarni, C. E. Jelden

The kinetics of precipitation of copper on iron of various purity were carried out under controlled conditions. The rate of reduction has been correlated with such parameters as copper and hydrogen ion concentration, geometric factors, flow rate, and temperature. The character of the precipitated copper as a function of flow conditions and rate of PreciPitation has been observed under a variety of conditions. ThE precipitation of copper in solution by cementation on a more electropositive metal has been known for many years. Basile valentine&apos; who wrote Currus Triumphalis Antimonii about 1500, refers to this method for extraction of copper. Paracelsus the Great2 who was born about 1493 cites the use of iron to prepare Venus (copper) by the "rustics of Hungary" in the "Book Concerning the Tincture of the Philosophers". Agricola3 in his work on minerals (1546) tells of a peculiar water which is drawn from a shaft near Schmölnitz in Hungary, that erodes iron and turns it into copper. In 1670, a concession is recorded4 as having been granted for the recovery of copper from the mine waters at Rio Tinto in Spain, presumably by precipitation with iron. Much has been published in recent literature on the recovery of copper by cementation, the majority of the articles being on plant practice.5-24 The rest include articles on investigation of the variables involved25-28 and a review of hydrometallurgical copper extraction methods." This literature has established: a) The three principal reactions in the cementation of copper are Cu + Fe — Fe+4 +Cu [ 11 One pound of copper is precipitated by 0.88 lb of iron stoichiometrically. In actual practice about 1.5 to 2.5 lb of iron are consumed. 2Fe+3 + Fe — 3Fe+2 [21 Fe +2H&apos;-Fe+2 + H2 [3] Reactions [2] and [3] are responsible for the consumption of excess iron. Wartman and Roberson&apos;28 have established that Reactions [ I] and [2] are concurrent and much faster than Reaction [3]. b) Acidity control is important in the control of hydrolysis and the excessive consumption of iron. he commercial workable range is approximately from pH = 1.8 to 3." c) Iron consumption is closely related to the amount of ferric iron in solution. Jacobi" reports that, by leaving the pregnant mine waters in contact wi th lump pyrrhotite (Fe7S8) for 3 hr, all the iron was reduced to the bivalent condition and scrap iron consumption was cut to 1.25 lb scrap per pound of copper precipitated. He also reported that SO2 has been used successfully to reduce ferric iron to the ferrous state. d) The ideal precipitant is one that offers a large exposed area and is relatively free of rust. e) High velocities and agitation show a beneficial effect upon the rate of precipitation, as it tends to displace the layer of barren solution adjacent to the iron and also dislodges hydrogen bubbles and precipitated copper to expose new surfaces. Little work, however, has been published on the reaction kinetics of copper precipitation on iron. Cent-nerszwer and Heller20 investigated the precipitation of metallic cations in solutions on zinc plates. They found the cementation reaction to be a first-order reaction. The rate constant was independent of stirring for high stirring rates and they concluded that the rate is governed by a diffusional process at low stirring speeds and by a "chemical" process at higher stirring speeds where the rate reaches a constant value. This conclusion has been challenged by King and Burger30 who could not find any region where the rate was independent of the stirring speed, although the rate constant they had obtained for high stirring speed was greater than the maximum value of the rate constant reported by Centnerszwer and Heller (by a factor of six). King and Burger, therefore, concluded that the rate of displacement of copper was controlled only by diffusion. Cementation of various cations on zinc has been summarized by Engfelder.31 APPARATUS A three-necked distillation flask of 2 000-mm capacity was used as a reaction vessel. A pipet of 10-mm capacity was introduced through one of- the side necks, the sample of sheet iron, mounted in a rigid sample holder, through the other, the stirrer being in the middle as shown in Fig. 1. The whole assembly was immersed in a constant-temperature bath. The stirrer was always placed at the same depth in the solution. EXPERIMENTAL PROCEDURE Reagent-grade cupric sulfate (J. T. Baker Chemical Co., N.J.) was used to make up a stock solution containing 10 g of copper per liter which was then diluted to various concentrations as required. Experimental data were obtained by measuring the amount of copper and iron ions in solution at successive time intervals. The initial volume of the solution was always 2000 ml, 10-ml aliquots being removed each time for chemical analysis. Because the total volume change of the solution was less than 10 pct, no correction was used for solution volume change. Nitrogen was bubbled through the solution before and

Jan 1, 1968

By D. Kuhlmann-Wilsdorf

A new theory of work hardening is developed which rests on only a few simple principles and is applicable to a wide variety of materials and dislocation structures. It explains, qualitatively, the general characteristics of the three-stage shear stress-shear strain curve of fcc metals, as well as the most important features connected with easy glide. The value of, the work-hardening coefficient in stage 11, is derived quantitatively, from first principles, and is shown to be insensitive to many changes in detailed dislocation behavior. The low rate at which energy is stored during mechanical working in stage 11, the dependence of slip line length on stress, and the so-called Cottrell-Stokes law are explained. Although the theory is primarily developed for fcc metals and alloys, it is largely applicable also to some other materiuls, and in particular apparently to poly crystalline simple steels in theh linear range of hardening. FOR pure fcc metals and alloys, after many different pretreatments and under a great variety of testing conditions, the well-known three-stage work-hardening curve is commonly observed. Moreover, the work-hardening coefficient in stage 11, 011, bears an almost constant relationship to G, the modulus of rigidity, such that K = G/Brr" = 300, within a factor of about 2 either way. The very fact that the three-stage curve and the value of K are so very persistent, can hardly be understood except on the assumption that some quite simple phenomena are responsible. In this light all available theories of work hardening must be judged unsatisfactory: Each of them is based on some quite specific dislocation model, while it is known that metals with similar values of K may have widely different dislocation structures, for example, dislocation tangles in the case of pure fcc metals, and piled-up groups of dislocations in a-brasstype alloys. Thus, while a particular theory might conceivably represent a special case with a reasonable degree of accuracy, it cannot possibly have illuminated the underlying principle. Another severe criticism applying to the presently most widely discussed theories is that they employ empirical or even quite unexplained parameters in order to arrive at a numerical value of K. This indicates that some vital aspect of the work-hardening process has remained unexplained. In the present paper, a new theory of work-hardening is developed which does not suffer from the above shortcomings. It is derived from first principles, and is founded on the realization that the underlying causes for the three-stage work-harden- ing curve and for the persistence of the numerical value of O,, must be few and simple. As a first step it is shown herein that the experimental oklservations on "easy glide" can be explained on the assumption that in this stage dislocations multiply, beginning in a number of restricted area:;, and from there penetrate into crystal regions still substantially free of glide dislocations until a quasi-uniform dislocation distribution has been established. During "easy glide", the resistance to the movement of the foremost dislocations, advancing into still largely a dislocation-free regions, is believed to control the flow stress. This resistance is determined by various factors, among them the reaction of the dislocations against bowing-out into loops, which is due to their line tension. Without such bowing-out no dislocation multiplication would occur, but the corresponding contribution to the flow stress stays practically constant all through stage I. According to the present theory, stage TI begins as soon as there are no more areas left into which newly formed dislocations have not yet penetrated, i.e., as soon :is a quasi-uniform dislocation density has been attaned. Regardless of how, in detail, the dislocations should happen to be arranged at the end of "easy glide", whether in pileups or tangles, on one glide system or on several, the stress necessary to overcome the line tension of the segments must now increase, because the dislocation density increiises while the average free dislocation length decreases. It is argued that all other contributions to the flow stress stay at about the same level as in stage I, and that the said increase in the stress necessary to bow out dislocation segments into loops is responsible for the major part of work hardening in stage 11. Stage TII of the stress-strain curve is explained by largely follouing the ideas of Seeger and his co-workers; however, it is pointed out that the start of stage m may on occasion indicate the onset of "conservative climb" instead of cross slip. Indirect evidence indicates that linear work hardening persists even under the profuse action of cross slip, but at a lower rate. Only when, in addition to cross slip, climb operates extensively, does the work-hardening rat,. drop sharply, ultimately down to zero. BRIEF SURVEY OF EXPERIMENTAL EVIDENCE The stress-strain curve of crystals of many different substances are qualitatively similar. No significant permanent plastic deformation takes place below the so-called "critical resolved shear stress". As the applied stress is raised beyond this level. yielding occurs with little or no work hardening and

Jan 1, 1962

By A. J. Chellew, Robert Condit

THE Company maintains a port agency at Antofagasta for dispatching copper and also for receiving from ocean vessels the vast variety of materials and supplies required at the plant. Copper shipments from the oxide plant refinery to Antofagasta amounted to 162,000 metric tons during 1951, while incoming freight totalled 175,000 metric tons. Electrolytic market copper in various shapes is currently loaded at the plant upon 20 or 30-ton cars of the meter-gage Antofagasta-Bolivia Railroad, to be transported a distance of 165 miles to the company&apos;s yard at Antofagasta. Here it is unloaded by cranes at the rate of about 175 tons per hr. Copper not scheduled for immediate shipment is stocked in tiers within the yard, but most of it is reloaded and sent to port works or a wharf serving lighters. Some of the vessels entering Antofagasta tie up at the port works, while others anchor inside the breakwater. In the latter case, all of the loading into the hatches is from 50-ton lighters towed to the vessel. When the vessel is at the dock, about half of the copper is loaded by port works cranes and the other half is loaded by ship&apos;s tackle from lighters. Fuel Oil Requirement Expands One of the innovations caused by the construction of the sulphide plant has been the necessity to provide facilities at Antofagasta for receiving and storing greater quantities of fuel oil than in recent years. From here it is loaded into tank cars for transportation to Chuquicamata. The older installation consisted of a 10-in. submarine pipe line, heating, and pumping facilities, a 67,500-bbl storage tank, and railroad loading rack, located at Antofagasta&apos;s northern limits. Two new tanks, of 80,000 and 30,000-bbl capacity respectively, have now been added and a second submarine pipe line will shortly be installed. Fuel oil requirements for the sulphide plant are estimated at 39,000 bbl per month, in addition to which 9000 bbl per month are required for the operation of the oxide plant. The company has a special lighter which is used by its diver for connecting the oil line with the oil tanker when disembarking oil, this lighter also being used for inspecting and overhauling buoys, as well as for recovery of bars of copper which are sometimes dropped into the bay when loading vessels during periods of rough weather. Transport of oil to Chuquicamata is by 76 tank cars of, 20-ton capacity and 20 of 30-ton capacity, all assigned to the company&apos;s fuel oil shipments excepting six for Diesel oil and one for gasoline. Incoming Freight Other freight received by the port agency has been averaging about 4800 metric tons per month of operating supplies for the oxide plant (renewals of equipment, spare parts, lubricating oils, and greases, etc.), some 1900 metric tons per month for the company stores, and about 6000 metric tons per month of materials and equipment for the construction of the sulphide plant. The greater part of these shipments (particularly the operating and construction supplies) originate from abroad, and must be handled through the Chilean Customs. Hence the company maintains a Customs storage warehouse (goods pending appraisal of Customs duties) in addition to its other storage facilities at Antofagasta. Not the least of the Port Agency&apos;s duties is to meet air and steamer passengers en route to Chuquicamata and to arrange for their trips up the hill. Landing in a strange country with a strange language, the first impressions of a new employee might otherwise be indeed confusing when facing Antofagasta&apos;s unique setting of brown barren hills.

Jan 1, 1952

By Hisaaki Takata, Nobukazu Nanko, Ichiro Izawa

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

Jan 1, 1981