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By J. H. Westbrook

UNTIL very recently the development of high temperature alloys has been strictly empirical. It is, in fact, a great tribute to the intuition, perseverance, and industry of the practicing metallurgists of this country and abroad that the amazing increases in gas turbine service temperature over the past 15 years have been achieved with so little fundamental knowledge. The result of this development has been the realization of successful but highly complex alloys possessing five, six, or more alloying elements and correspondingly complex microstructures. Within the last few years, considerable effort has been devoted to the identification of the secondary phases in these alloys and to the correlation of the presence or absence of such phases with alloy properties. Little attention, however, has been given the problem of identification of the mechanism (s) by which such phases enhance the properties of the alloy and especially to the assessment of the particular properties of the secondary phase (either in- trinsic or relative to the matrix phase) which are basically responsible for the improvement. Conceivable mechanisms which might apply include dispersion hardening and precipitation hardening. In both cases geometrical factors—particle size, dispersion, volume fraction, etc.—and agglomeration re-

Jan 1, 1958

By G. D. Bates

INTRODUCTION The Magmont mine is a joint venture of Cominco American Inc. (operator) and Dresser Minerals, Inc. The mine-mill operation is located approximately 160 km (100 miles) southwest of St. Louis, MO, in what is commonly referred to as the "Viburnum Trend.” The Magmont mine is designed for a production rate of 3810.2 t/d (4200 stpd) on a 5-day week, three shifts per day basis. Initial production began in 1968. The mine is open stope, room-and-pillar, and essentially horizontal along the trend of the ore body. Briefly, the main geological structure can be described as a brecciated graben bounded by reverse faults. The ore body in cross section is shaped like a bell curve with some lateral extension at the lower part. Presently outlined ore is 609.6 to 762 m (2000 to 2500 ft) in width and 2133.6 m (7000 ft) in length. The ore varies in thickness from 4.87 m (16 ft) on the fringes to an average of 27 m (90 ft) in the high ore areas bounded by the reverse faults. Lead is the primary metal with zinc and copper secondary. MINE DESIGN The basic design of open stope, room-and-pillar mines has been described by several writers and need not be repeated here. (Anon., 1970; Bullock, 1973; Casteel, 1972; Christiansen et a]., 1970; and Lane, 1964) This discussion covers the mining sequence as applied to the particular problems at the Magmont mine, the use of equipment, and deployment of the work force. In the upper portion of the Magmont ore body is a layer locally called the False Davis shale. This layer lies below the true Davis shale, is normally interbedded with dolomite, is of varying thickness, and if mineralized, is included in the top pass of the mining sequence. In thick ore areas this layer will be 2.13 to 2.43 m (7 to 8 ft) in thickness and will occur in the upper portion of the pillars. Due to its incompetency the presence of this False Davis layer is of primary concern in mine planning and operation. Mining areas are divided into three basic groups by ore thickness. First are areas of ore up to 6.09 m (20 ft) in thickness. These areas are below the False Davis shale and are mined single pass with drill jumbo. Second are those areas up 13.71 to 15.24 m (45 to 50 ft) in height. The first 4.87-111 (16-ft) Pass is taken at the top of the ore and the back and pillars secured. Benching the lower portion(s) in 4.57 to 4.87-m (15 to 16-ft) passes is then done with either a drill jumbo drilling horizontally or a crawler drill drilling vertically. Normally these areas are below the Table 1. Productivities per Manshift False Davis shale. These areas may also be mined by back slashing, or overhand benching, where the first 4.87-m (16-ft) pass is taken at the base of the ore and successive 4.87- m (16-ft) passes are taken upward. A minimum amount of back slashing is done at Magmont since it requires repetition of roof control on each pass and roof control is the single largest stoping cost at Magmont. Ore left to provide a working platform oxidizes and is coated by oil spills thus reducing metallurgical recoveries. The third mining area is over 15.24 m (50 ft) in height UP to a maximum of 40.23 m (132 ft) and will encompass the False Davis shale. These areas are mined by first driving +15% inclines to the top of the ore body. The top pass is mined and the back is bolted and roof mats installed as a matter of standard practice to minimize roof problems as mining progresses downward. Once the back and pillars on the top pass are secured, benching begins on successive passes with either the drill jumbo or crawler drill. Pillars on all successive passes below the top pass are secured as necessary. While benching progresses below the top pass, the pass at the base of the ore body is mined leaving a sill of 4.57 to 7.62 m (15 to 25 ft) in thickness to be removed with the crawler drill in a retreating manner. Rooms are mined on a 1.57 rad (90") grid pattern to insure alignment of pillars where multiple passes are taken. Pillars are designed on a 17.98-m (59-ft) spacing with rooms up to 10.66 m (35 ft) in width. Heading widths are wide enough for the mobile equipment to turn without additional allowance for curves. The result is a flexible layout which provides a maximum number of headings available for high extraction rates and grade control. PRODUCTION Incentive Bonus Incentive bonuses play an important part in the mine production at Magmont. Production crews are trained to perform only one of the mining functions of drilling, blasting, mucking. or roof bolting. This specialization, or functionalization, is augmented by development to open all possible stoping areas as early as possible in the life of the mine. This insures that each crew will have enough headings to perform its specialty. The incentive bonuses increase exponentially as output increases. The lucrative incentive bonus coupled with the specialization of the production crews and proper mine development have combined to give the high productivities shown in Table 1. Development crews perform all mining functions in their area. The incentive bonus is paid on a per foot basis, Crews on different shifts working the same heading share equally in the bonus proportional to their contract hours worked.

Jan 1, 1982

By A. López-Valdivieso, M. A. Corona-Arroyo, M. Llamas-Bueno

Magnetite (Fe3O4) ore is concentrated by low-magnetic-field drums to recover a magnetite concentrate that is low in silica (SiO2). The presence of SiO2 in the magnetite concentrate for steel production increases the steel processing costs, so a major challenge in magnetic concentration is to lower the SiO2 grade in the Fe concentrate. This work presents studies that were carried out on the removal of SiO2 from the magnetite concentration at the plant scale. Studies were performed with a three 36 × 96″ drum unit processing 46 ton/h of rougher magnetite concentrate. These studies showed that silica appears in the magnetite concentrate by three mechanisms, namely SiO2 entrainment in Fe3O4 chains, heterocoagulation between Fe3O4 and SiO2 particles, and mineral locking of SiO2 to Fe3O4. The SiO2 entrainment mechanism had the highest contribution (75%) to the SiO2 recovered in the Fe concentrate. Electrokinetic studies on Fe3O4 and SiO2 showed that heterocoagulation occurs because of the low negative zeta potential of the minerals at the pH of the plant slurry. This study also showed that a high percentage of ultrafine size SiO2 particles trapped in Fe3O4 agglomerates was not removed by the dilution water for the slurry fed to the drum. To lower the SiO2 recovery in the drum magnetic concentration process, work should be directed towards the removal of trapped SiO2 in Fe3O4 agglomerates.

By R. D. Lrwin, W. M. Mullins, S. Venugopa, J. C. Malas

"Physical models for acoustic emission (AE) are introduced and expressions are derived to predict AE activity from such parameters as applied stress, strain and strain-rate. These models are then incorporated into a visco-plastic finite-element simulation program, and the acoustic emission event rate generated during metal forming operations are predicted. Simulation results are presented for upsetting operations on a typical C-Mn type steel for various friction and die geometry conditions. The AE model predictions compare well with the limited available experimental results reported in the literature and demonstrate that AE signatures can be reliably simulated. The signatures predicted for metal forging are fairly insensitive to processing parameter changes. This suggests that AE event rate monitoring may not be well suited to process monitoring for forging. However, the size of the features is conjectured to influence the frequency content of the AE signals. Thus, since current simulations include only the ability to predict the number of events for a set of processing conditions, future work is required to simulate the duration and magnitude of acoustic events. Fourier analysis techniques may then be applied to investigate correlation with microstructural features.IntroductionThe open loop control system method for deformation processing design, as described by Malas et al. [1 ], consists of several steps in series to achieve the final design. The first step uses microstructure development models to determine the optimum deformation path to achieve the desired final microstructure. The second step, the process model, calculates the processing parameters required to achieve this deformation path. Finally, the metal-working system uses the required processing parameters to produce the part. Figure 1 shows this in block diagram form"

Jan 1, 1997

By Scott Jones, Tony R. A. Durkac, David W. Stephens

Williams Mine was considering spending a large amount of capital to install a pumping system underground to rid the mine of slimes; with the introduction of Delvo Stabilizer this challenge would be met. This paper will address the process to introducing the Delvo Technology into the Cemented Roddill Program. Some areas that were worked on included: testing of the mix design, bet dosage rate for an economical solution, temperatures at different levels of the mine, the amount of time needed once the grout was put to sleep, preventing plugs in the delivery system A controlled dosage that gave the Operator a 12 hour window of flowable material was developed and an added feature to the introduction of Delvo was also noted; an increase in MPa strength over the design. Collectively with the help of operators at the batch plant and the underground miners involved with the program all concerns where addressed and all goals met. A cost effective program was implemented thus increasing productivity, preventing costly plugs, and improved the actual maintenance on the entire delivery system, which included the batch tanks, aggitator tanks, all pumps required in the system and of course reducing buildup of slimes at the cones.

Jan 1, 2001

By E. J. Lintner

This paper is one of a series being prepared by the United States Bureau of Mines describing clay mining and crushing methods and costs at various operations throughout the United States. These papers are designed to disseminate technical information regarding methods used. The cost tabulations represent local operating expenditures only and not total production costs. It is recognized that publication of total production costs might perhaps, in many instances, cause embarrassment to the individual operator as well as to the industry as a whole. On the other hand, operating costs are essential to the technical discussion and study of the methods employed. The attention of the reader is specifically called to this distinction in order that no misunderstanding of the scope of the cost tabulation shall ensue. The following report deals specifically with the methods employed by the Camp Brothers Company, Mogadore, Summit County, Ohio.

Jan 1, 1936

By Sunil Katharia

In February 2016, the upgraded Blast Furnace 1 of JSW Steel Limited (JSW) at its Vijayanagar Steel Plant was blown in after a 6 months shutdown in which the whole furnace was rebuilt on the existing foundation and within the existing tower structure. Despite several plant constraints it was possible to increase the furnace inner volume by 85% - reaching 2307 m3 from existing 1250m3. The upgradation project covered the stockhouse and its dedusting system, top charging system, furnace profile, furnace cooling system, casthouse and its dedusting system, slag granulation plant, pulverized coal injection plant, hot blast system and gas cleaning plant – with major interventions in order to cope up with the changed operating conditions. The total scope of the project was split by JSW into several packages, all based on design data provided by Paul Wurth. JSW was responsible for coordinating the local subcontractors along with Paul Wurth which has been appointed as main technology supplier in charge of complete basic and selected detail engineering in the key areas and of the supply of core equipment such as the Bell-less Top®, copper staves, GCP key items and TMT casthouse machines. The main design concept and the key points of the project such as optimization of project ROI, advanced erection concepts as well as a summary of the operating results are presented confirming JSW “BF1” upgradation as one latest examples of the current state of the art in the revamping projects.

Aug 17, 2017

By D. R. Tesarik

The factor of safety, the traditional measure of stability for earth embankments, is graphically presented in factor of safety charts in this Bureau of Mines report. Factor of safety contours are drawn for homogeneous earth embankments for two contrasting situations: (1) no phreatic surface and (2) a phreatic surface that assumes 10-pct freeboard. The factor of safety can be read directly from the charts if the physical properties of the soil and the geometry of the embankment are known. The curves provide a quick first approximation of the "most stable" and "least stable" condition of the embankment. The Simplified Bishop Method was used for the safety factor calculations, and a sequence of least squares curve fitting steps was used to present the factor of safety in final chart form.

Jan 1, 1981

By Giovanni Barla

The stress-relief method for the determination of the ground stress tensor is a common procedure in rock mechanics. In particular, the method of applying gauges to the end of a borehole is a well-established technique. Investigations, both experimental and theoretical, indicate that large discrepancies can exist in the determination of the ground stress tensor when the rock substance is not the ideal elastically isotropic material assumed for the reduction of data. The purpose of this paper is to extend the analysis of strain gauges applied to the end of a borehole to rock material more closely approximate to an orthotropic substance.

Jan 1, 1969

Attempt has been made to determine the nature of charge on activated carbon surface by studying the electro-kinetic properties of the carbon in the presence of cations, anions and aurocyanide solutions. The zeta-potential measurements were made in the presence of individual ions and also in combinations. The cations and anions chosen were those commonly associated with gold bearing ores in Kalgoorlie area. It was found that the iso-electric point of Norit Carbon was at pH 4.45; below this pH the charge was positive, while above, it was negative. In the absence of gold cyanide solutions and in the presence of individual cations the i.e.p. shifted to the alkaline range (pH 7.2-8.5) while in the presence of anions the i.e.p. was shifted to less than pH 3.5. In the presence of aurocyanide solutions, the iso-electric point was still in the acid region irrespective of whether anions or cations were present. The probable mechanism for such a behaviour has been attempted after measuring the specific conductivities of solutions before and after gold adsorption, by studies using the electron- probe-mirco-analyser, and S.E.M. on the carbon surface.

Jan 1, 1982

By E. F. Adkins, R. W. Douglas

The variation of the spectral emissivity of osmium and ruthenium with temperature can be expressed by the following relations: DURING a study of the sintering characteristics of osmium and ruthenium, discrepancies in the previously accepted melting points of the metals were noted. The melting temperatures had not been measured at the time of the work, but were estimated in the literature as 2700°C for osmium and 2450°C for ruthenium.' A recent investigation by Baird2 has shown the melting temperatures of osmium and ruthenium are 3000" and 2250° C, respectively. In the present work, osmium bars sintered at temperatures as high as 2900° C showed no evidence of melting, whereas ruthenium showed definite evidence of melting at temperatures as low as 2250°C. In view of the apparent discrepancies, the melting temperatures of both metals were determined. The spectral emissivities of both metals were determined as an aid to studies of the sintering characteristics of these metals. MATERIALS AND PROCEDURES The materials used in these studies were metal powders supplied by Englehard Industries, Inc., through the courtesy of The International Nickel Co. Analytical data from three laboratories, although not in complete agreement as to the purity of the material or the major impurity elements, indicated that the maximum impurity levels were 0.5 and 1.0 wt pct in osmium and ruthenium, respectively. The metal powders were pressed to 1/4 by 1/4 by 6-in. bars at 20Tsi using carbon tetrachloride as a lubricant. The pressed bars were presintered 1 hr at 1200°C after heating 1 hr at 400°C. The presintered densities of the several bars of the metals ranged from 65 to 80 pct of ideal density. Specimens were self-resistance heated in a vacuum sintering bell. Temperatures were measured with an L&N, disappearing filament optical pyrometer. True temperatures were read at the base of an 0.30-in.-diam, 0.150-in. deep hole drilled into the center of a presintered bar. Apparent temperatures were read from the surface of the bar adjacent to the hole. MELTING POINT DETERMINATION The melting temperature of the metals was determined as the temperature at which a slight additional increase in current through the specimen resulted in a reduced temperature at the base of the hole. The apparent reduction in temperature is a result of the liquid metal filling the hole, thus altering the black-body conditions. The melting points of osmium and ruthenium, based on three deter-

Jan 1, 1962

By M. Kosugi, M. Kawata, J. V. Satur, K. Mitsuhashi, R. Kawarabuki, W. Kracht, Y. Ebisu

Copper concentration is becoming more difficult because remaining copper ores are more complex and of finer copper mineral grain size, making copper minerals harder to float because of the fine particle size required to liberate the valuable species. The purpose of this work was to study the flotation behavior of fine copper sulfide ores at different conditions of agitation and aeration. The experimental campaign was carried out in a laboratory Denvertype flotation cell of 1 L volume modified to be operated with forced air that allows decoupling air flow rate and agitation speed. Four ore samples from Chile were used, all of them containing copper sulfides such as chalcopyrite, bornite, chalcocite, and tennantite, but with varying grain sizes and compositions. Preliminary flotation tests for establishing the flotation conditions of grinding size and collector dosage were conducted before the kinetic flotation tests that were carried out at different agitation speeds varying from 900 to 1300 rpm for gas flow rate ranging from 1.6 to 7.2 L/min. The flotation products were analyzed by ICP-MS and MLA (Mineral Liberation Analyzer). The results showed that high agitation speeds with corresponding air flow rates resulted in better recovery than moderate agitation speed and air flow rate, and an excess in gas flow rate resulted in bad performances. The flotation rates of the fine copper minerals were improved despite the different improvement rates among the fine size fractions. The impact of these operational variables, however, was different when comparing the results for the four ore samples. While it was found that the order of floatability in the four samples was the same, that is, chalcocite > tennantite > bornite > chalcopyrite, the kinetic constants of the copper minerals in the samples were different, indicating that mineral composition in the sample affects the flotation kinetics.

Jan 1, 2018

By S. R. Seagle, R. Schuhmann, N. A. Parlee

Rates of CO evolution and CO absorption were measured for liquid-iron alloys containing from 0.15 to 4.4 pet C, using a modified Sieverts apparatus. The alloys were held in alumina crucibles, so that both crucible-metal and gas-metal reactions occurred simultaneously. The data are interpreted on the hypothesis that the C-O reaction rate is controlled by oxygen diffusion across the boundary layers at the gas-metal and crucible-metal surfaces. CARBON-monoxide evolution from liquid Fe-C-0 alloys is a key reaction in steelmaking processes, both in the steelmaking furnace and in the ingot mold. Also, this reaction probably is an essential step in the desulfurization of iron under blast-furnace conditions, for which carbon is the principal reducing agent or deoxidizer present in the iron. The thermodynamic properties of the solutions of C and 0 in Fe and the equilibria of these solutions with gaseous CO and CO, have been investigated in some detail and are reasonably well understood. However, while these equilibrium data have accumulated in the laboratory, other data have accumulated to show that equilibrium conditions are often not achieved under plant operating conditions. Thus, an understanding of the rate and mechanism of the carbon-oxygen reaction not only has theoretical interest, but ultimately may assume considerable practical importance. Reaction Mechanisms and Rate Equations The over-all reaction under consideration is C (in liquid Fe) + 0 (in liquid Fe) + CO (gas) [I] Although this appears to be a straightforward heterogeneous reaction, involving just two phases (liquid metal and gas), a wide variety of reaction mechanisms has been proposed. Different authors have expressed divergent views as to the nature of the rate-controlling steps. Accordingly, a brief discussion of the possible reaction steps will be given. 1) Homogeneous Reaction Within Liquid Iron— Feild' and Jette2 have applied chemical reaction-rate theory as if the C-0 reaction were a homogeneous, second-order reaction occurring within the liquid-metal phase. This mechanism necessarily produces CO "molecules" dissolved in liquid iron. In order to yield CO gas, the dissolved CO must nucleate CO gas bubbles, and the dissolved CO also must be transported to the bubble surface where it enters the gas phase. However, this mechanism is inconsistent with present concepts of the structure of liquid-metal solutions. 2) Reaction at Gas-Metal Interface—At the surface of contact between gas and metal, a second-order reaction between dissolved C and 0 may occur to produce a CO molecule which enters the gas phase. This reaction presumably involves an "activated complex" structure in the interface. Some such surface mechanism appears essential because the carbon and oxygen do not occur in the same species in the two phases and therefore must react in some way at the surface. At liquid-iron temperatures, calculations based on reaction-rate theory indicate that the surface reaction must proceed extremely rapidly and thus cannot be a rate-determining step in the C-O reaction under ordinary conditions.3 ccordingly when the over-all reaction is measurably slow, the surface reaction may be considered at equilibrium; that is Cc* Co* = m'Pco [2] in which Co* and Co* are the concentrations of carbon and oxygen (weight per unit volume of metal), respectively, in the liquid-metal phase at the interface, Poo* is partial pressure of CO in the gas phase at the surface (atmospheres), and m' is the mass-action constant for Reaction [I]. Since activities of C and 0 are not strictly proportional to concentrations, m' is not a true thermodynamic equilibrium constant but varies with carbon content.' 3) Mass Transport—If the reaction is to proceed at the gas-metal surface, as described above, dissolved carbon and oxygen in the liquid iron must come to the surface and gaseous carbon monoxide must move away from the surface. Details of the mass-transport mechanisms depend on the kind of system under consideration. For a stirred metal bath in contact with a CO gas phase, the transport of C and 0 to the surface can be described by 1 Do Mols CO evolved per sec = — 1/16— Do/do Agm (CO-Co*) = 1/12 Do/do Agm *(Co-Co*) [3] In this equation, Do and D, are the diffusion constants for oxygen and carbon in liquid iron, 6, and 6,. are the boundary-layer thicknesses, A,,,, is the

Jan 1, 1959

By P. K. Foster, A. McNabb

A consideration of the literature shows the inadequacy of Fick's Laws to describe the diffusion of hydrogen in steels below 400°C. These laws are modified an the assumption that hydrogen is delayed at fixed sites in the lattice. Approximate and exact solutions of the resulting nonlinear system are derived for simple geometrical shapes. The results qualitatively account for the observed anomalous behavior. Experimental methods for verifying the theory are suggested. 1 HIS paper presents a new analysis of the diffusion behavior of hydrogen in steel based on a physical model which supposes that hydrogen atoms wander in a random manner through the crystal lattice, but tend to get trapped or delayed at certain fixed sites uniformly distributed throughout the metal. These sites, which will be called traps, are regarded merely as potential wells of significantly greater depth than those encountered in regular regions of the crystal lattices and no assumptions are made concerning their nature or origin. The results of experimental studies of the diffusion of hydrogen in iron and steel have usually been reported in the literature in terms of a diffusion constant D derived under the assumptions that the hydrogen concentration C is governed by Fick's laws.1 Fig. I is a composite of published graphs of the man. ner in which the values of D, so derived, vary with temperature. It is convenient to consider this material in two categories; the high temperature results above 400°C, and the results at lower temperatures. The graphs 2, 3, 4, 5 in the first class lie within a factor of three of each other and are relatively consistent compared with the low-temperature values. There is an increasing divergence in values of D with decreasing temperature till at room temperature they are spread over three orders of magnitude. Small values of D have been obtained for work-hardened specimens.' It is evident that D is a function of other variables besides temperature in this range and that these neglected variables are in some way connected with the work hardening experienced by the specimen. In view of evidence for an interaction between hydrogen in solution and dislocations9'10 and adsorption on crack surfaces,' some doubts must be entertained about the validity of Fick's laws and their associated physical model of noninteracting particles in random motion through the medium. More direct evidence of their inapplicability to the diffusion of hydrogen in steel was first produced by Darken and smith." Consider the permeation of hydrogen through a plate of thickness a under a boundary concentration Co at the ingoing, and zero at the outgoing, face. If Fick's first law1 is valid, the steady state flux P, through unit area is DCo/a. In addition, the concentration C varies linearly throughout and Qo, the mean quantity of hydrogen in the medium, is co/2. Compare Qo with Q1 the mean quantity required to saturate the same specimen to a uniform concentration G. In this case Ql = Co and hence the ratio

Jan 1, 1963

By Mark W. Masson, David B. Stevenson, David W. Broughton, Douglas R. Cruji, Steven E. Parry

"The Amalgamated Kirkland gold deposit is located at Kirkland Lake, Ontario, within the Archean Abitibi Greenstone belt of the Canadian Shield. It lies on a fault or “break”, termed the Cyprus Break, which is a splay off a regional structure termed the Cyprus-Battle Mountain Break. The Cyprus Break merges with the Cyprus – Battle Mountain Break above the -250 m level, as it was not encountered in shallow drilling or surface exploration. These structures lie 1.5 km south of the 727 000 kg (23 million ounce) Kirkland Lake Main Break and 0.5 km north of the 872 000 kg(30 million ounce) Larder Lake Break.The deposit is hosted by alkalic Timiskaming Group volcanic and sedimentary rocks which are overthrust by older Larder Lake Group tholeiitic and komatiitic fows along the Larder Lake Break. Timiskaming syenitic intrusives, similar to the host rocks for the Kirkland Lake Main Break gold deposit, intrude both suites.The Amalgamated Kirkland deposit consists of lode-style gold mineralization which is mainly hosted by altered and pyritized Timiskaming trachytic volcanics which wedge out or thin at depth, between two sedimentary units. The intersection of the Cyprus – Battle Mountain and Cyprus Breaks, in conjunction with the westward plunge of the volcanic wedge, appears to control the westerly plunge of the Amalgamated Kirkland deposit at approximately 50º.Mineralization is characterized by blue-gray brecciated and “wormy” quartz-ankerite veinswhich contain up to 10% fne-grained pyrite and, in decreasing order of abundance, minor galena,chalcopyrite, sphalerite, molybdenite and visible gold. The sulfdes and gold commonly occur alongfractures and wallrock inclusions within the veins. Native gold occurs as fne pinpoints distributed in 1-mm to 5-mm size clusters of up to ten or more grains.The auriferous veins occur within a quartz-ankerite-sericite-pyrite alteration assemblage which is enveloped by a broader zone of ankerite-sericite±hematite±quartz-pyrite alteration up to 60 m wide. This latter assemblage overprints the regional greenschist-facies chlorite-calcite assemblage.The Amalgamated Kirkland property has had a long history of exploration activity beginning with the first discovery of mineralization in 1911 on the Hunton Shaft claim located in the northeast corner of the property. Since that time various prospecting, mapping, trenching, geophysical surveys and diamond drilling programs have concentrated on either carbonate zones along the Larder Lake Break or quartz veins along the northern side of the property in association with complex trachyte flows and syenite dikes.Battle Mountain Canada Inc. located the Cyprus – Battle Mountain Break through a prospecting, trenching and geophysical program in 1989. Drilling by Battle Mountain and Cyprus Canada under the surface showing, coupled with several revisions to the structural interpretation of the Cyprus –Battle Mountain and Cyprus Breaks, resulted in delineation of the Amalgamated Kirkland deposit. Location of an auriferous structure through basic surface exploration within 1 km of a world-class orebody demonstrates both the subtle nature of Archean gold deposits and the potential for finding new mineralization in heavily prospected camps."

Jan 1, 1995

By W. C. Winegard, E. L. Holmes

The effects of the solute elements lead and bismuth on grain growth in zone -refined tin have been investigated. At low temperatures, viz. below 160°C. or at high concentrations, viz. greater than 0.03 at. pct, the temperature dependence of grain growth was found to be constant. A correlation was found to exist between the rates of grain growth in the presence of solute impurities and the factor [C Ir - 1]2 , where C is the solute concentration, r is the radius of solvent atom, and r' is radius of solute atom. It is shown that the Liicke and Detert theory does not give a satisfactory explanation of the results obtained. INVESTIGATIONS of grain growth in zone-refined lead and tin and some effects of solute elements on grain growth in these metals have been reported by Bolling and winegardly2 and Holmes and winegard? The present investigation is an extension of the previous work, and was undertaken to provide additional results which might make it possible to discover the mechanism whereby the rate of grain growth is affected by the presence of foreign solute atoms. The experimental procedure was exactly the same as that described previously.3'4 Figs. 1 and 2 show the results obtained with lead and bismuth additions over the same temperature range, viz. 167" to 220°C, as that used for the investigation of grain growth in zone-refined tin.3 Fig. 3 shows the results obtained for zone-refined tin and for this tin with lead solute additions over an extended temperature range. DISCUSSION OF RESULTS As mentioned previously,3 the character of grain growth in the particular zone-refined tin used in these experiments appears to change as the temperature of annealing falls below about 160°C. The grain growth became somewhat irregular and was characterized by a jaggedness of many of the grain boundaries. The reason for this change is not known, but may be due either to residual impurity effects or possibly an oxide precipitation. It was obvious that normal grain growth only occurred in the zone-refined tin through the high-temperature range 167" to 220°C, and so the effects of solute atoms on grai11 growth over the two temperature ranges will be discussed separately. a) Effect of Impurity on Grain Growth in the Temperature :Range Above 167°C. It is apparent from Figs. 1 and 2 that, for low concentrations of solute, the rates of grain growth tend towards a common value as the melting temperature of tin is approached. As the concentration of lead is increased, the rate of grain growth close to the melting temperature is also reduced, and at high concentrations the temperature dependence of the rate of grain growth appears approximately constant. It should be noteil that the effect of bismuth is greater than that of the same concentration of lead, but both these solutes have a smaller effect on the rate of grain growth than silver, as seen from an earlier investigation of Holmes and Winegard. It should

Jan 1, 1962

By Juhyung Lee

Recently, large diameter drilled shafts are often used for foundations of large-scale marine bridges with high capacity and required to resist lateral loads. Steel casings are essential in supporting the hole-diving concrete installation of the drill shafts constructed in the sea. The casings, however, are usually not considered as the structural elements and their load bearing capacity is often overlooked in the design of drilled shafts. As the diameter of drilled shafts becomes larger, there are higher demands for more efficient and economical design of large-scaled drill shafts in the sea. Steel-concrete composite columns are popular for the superstructure of bridges, and the outer steel attached to the shaft increases the shaft resistance through confining the concrete. In this study, the lateral resistance of steel-concrete composite drilled shafts was evaluated quantitatively based on the numerical analysis. When steel casings are used as structural elements such as the composite columns. Ultimate lateral bearing capacity of composite drilled shafts with various diameters was numerically calculated through the 3D finite element analysis. For that, elasto-plastic model with the perfect plasticity is involved to capture the ultimate load. A commercial FEA program, MIDAS-GTS, is used in this study. Real field conditions of the west coast, Korea were examined to set up the ground conditions and pile lengths required for this parametric studies. Detailed characteristics of the stress and displacement distributions are evaluated for the better understanding on the mechanisms of the composite shaft behavior.

Jan 1, 2008

By D. Duvris

The main objective of this work was to quantitatively investigate exhaust emission levels, as related to maintenance (service) intervals, during the operation of diesel-powered equipment at an underground bauxite mine in Greece. Such emissions depend on a great number of parameters, including the type (technology) and tuning of the engine, the type of work performed as related to the engine workload, suggested maintenance intervals, the altitude, the temperature and the humidity of the atmosphere of the underground mine, the properties of the fuel, the operator's training and experience level. Necessary ventilation quantities were calculated based on exhaust gas concentration measurements, site specific relationships between pollutants and adjusted air quality indices. Diagrams comparing necessary and recommended air flow for several equipment groups featuring two engine types and for different maintenance intervals are presented.

Jan 1, 1998

By P. Hazebroek, M. Prats, E. E. Allen

Well locations in water floods frequently do not coincide with the ideal well locations associated with the flood pattern which is used. A question of importance is. "How much oil is lost because of this?". This is the problem treated in this paper. It is assumed that each pattern in a five-spot flood has the proper number of production and injection wells but that the wells may be off-pattern. The distance by which a well is off-patern is given in terms of the pattern spacing, defined as the length of one side of a five-spot. The influence of off-pattern wells on the performance of a five-spot flood has been investigated mathematically for unit mobility ratio. All production and injection rates are assumed to remain equal. The loss in recovery resulting from pattern irregularitie: has been calculated at a cumulative water injection corresponding to a 98 per cent water-cut in a regular five-spot. Displacing a single production or injection well by about one-third the pattern distance reduces the total field recovery by 4.4/N per cent of the recovery obtainable with regular spacing, where N is the number of five-spots in the flood (assumed large). When off-pattern wells are separated by at least one regular five-spot and are displaced by less than one-third the pattern distance, the total loss in recovery in a large-scale flood is, at most, 1 per cent of the recovery obtainable from that field with a regular five-spot pattern. However, much higher losses may occur if the off-pattern wells are not separated by a regular five-spot, if the rates are imbalanced, or if the mobility of the injected water is higher than that of the crude. Although the loss in oil recovery due to pattern irregularity often may be low, the off-pattern case will always require injection of more water to reach the same final water-cut. The volume of water injected per unit volume of oil displaced has been calculated for each five-spot adjacent to a five-spot containing a single 08-pattern well. The cumulative water injection at the time of 98 per cent water-cut may be increased by as much as 37 per cent when a production well is displaced one-third the pattern distance toward an injection well. The existence of differences between actual subsurface locations and the idealized well patterns used in secondary recovery operations is well recognized. However, the methods commonly used for predicting waterflood behavior ignore such differences, even when subsurface locations are known. The reason for this is simple. The idealized well patterns are highly symmetrical and, as such, the production history of the pattern (and even of the field) can be determined from that of a few production wells—usually from only one. Consideration of the many irregular well spacings found in a specific field would entail a large amount of work, and then the results would be applicable only to that specific field. In an attempt to provide an answer to the more general problem, this paper considers a water flood having an idealized well pattern, except for one well which is not at its regular location. When several off-pattern wells are present, their contributions are assumed to be additive unless the off-pattern wells are close to each other. Only four positions of the off-center well (at x, y. of Fig. 1) are considered in this paper. These are shown

By H. T. Shore, J. M. Schultz

A number of experimental rnethods—X-ray powder diffractometry, Laue photography, X-ray small-angle scattering, and transmission electron microscopy and dijfraction—have been utilized to examine the morphology associated with precipitation from the terminal, g, solid solution of a Zn-1 pct Cu alloy. A significant age hardening was observed in a 1 pct Cu alloy. X-ray and electron diffraction results showed that the structural inhomogeneities associated with the hardening were isotructural with the matrix. The average size and shape of the inhomogeneities were deduced from the electron microscopy and X-ray small-angle scattering. The preprecipitates are hexagonal platelets some 300? in diam. and some twelve unit cells thick. The orientation of the platelets was deduced from Laue photographs and electron diffraction. The platelet plane is (0001). When a large amount of pre-precipitation is present in a localized volume the new lattice is often disoriented by a rotation about (0001) of of the matrix. WhILE dilute Zn-Cu alloys have been commercially important for some 50 years, relatively very little is known metallographically about this material. The "Zilloys", zinc with about 1 wt pct Cu and sometimes a small addition of magnesium, are used to produce rolled zinc which is harder and stronger than that produced by other rollable zinc alloys.' According to the phase diagrams of the zinc-rich side of the Cu-Zn system, such dilute Zn-Cu alloys should age-harden;2-5 the solubility of copper in zinc, g-phase, at 424°C is 2.68 pct, while at 0°C it is only to 0.3 pct. However, the published literature on the aging of this system appears to be limited to a documentation of the contraction of 1, 2, and 3 pct Cu alloys aging at 95°c,6 and an attempt to measure changes in lattice parameters during aging.' In the latter work, no lattice parameter changes were detected, although a broadening of the highest-angle lines was detected and considerable diffuse scattering was observed. Micro-structural investigations have been limited to the latest stage of aging, wherein Widmanstatten precipitates are formed.3,47 These alloys are of interest for still another reason. The two most zinc-rich phases in the Cu-Zn system, 77 and E, are both hcp. Moreover, the change in a, between 17 and t for a 1 wt pct Cu alloy is onlv 3.64 -,~ct: the change in Co is 12.0 ict. It would be anticipated that precipitation in such a material might occur through metastable phases or G.P. zones with epitaxy along mutual 0001 planes. The goals of the present work are aimed at partially filling the void of knowledge concerning the early stages of precipitation from the g phase. In particular, we have attempted to document the magnitude of the age hardening of this system and to determine the size, shape, and orientation within the matrix of the elements of precipitation in an early stage of condensation. EXPERIMENTAL A) Specimen Preparation. Specimens were prepared In two somewhat different ways, one method being used for X-ray Laue and diffractometer measurements, optical microscopy, and Rockwell hardness measurements and the other used for electron microscopy and X-ray small-angle scattering. In the first case zinc and copper in the proper proportions to yield a 1 wt pct Cu alloy were melted together in a closed graphite crucible. Castings so made were free of apparent segregation or oxidation. The castings were then solution-annealed at 400°C for several days and then quenched in water to room temperature. Filings of portions of the specimens were made for use as X-ray powder diffractometry specimens. The electron microscope material was made as follows. Castings were made under vacuum with copper powder placed inside a hollow zinc cylinder to insure good contact of the materials. These 1 wt pct Cu pieces were then rolled to 0.1 mm with an intermediate anneal in vacuo. The rolled sheets so formed were then annealed for about 6 hr at 225°C. Finally the specimens were electropolished slowly until thin enough for transmission electron microscopy. The polishing is discussed in greater detail in the Results section. B) Measurements. X-ray measurements of three types were performed. A G.E. XRD-5 diffractometer was used to examine powders of the alloy for identification of second-phase material. A Kratky small-angle camera, also operating from a G.E. tube, was used to investigate the sizes of small precipitate particles. In both cases, nickel-filtered copper radiation was utilized. Finally, individual grains of the large-grained castings were examined in the back-reflection Laue geometry. Electron microscope studies were carried out with a J.E.O.L. Model 6A instrument. RESULTS A) Hardness Measurements. Hardness measurements performed at room temperature on the large-grained polycrystalline specimens showed a hardening which was essentially complete in 3 hr. Fig. 1 shows a typical plot of hardness vs aging time. The relative magnitude of the ultimate hardening varied from run to run between 150 and 200 pct of the value for the material immediately after quenching from the solution anneal. Most probably the variations reflect small changes in the time taken to remove the specimen from the vacuum furnace after the solution anneal.

Jan 1, 1969