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By Roger M. Buckeridge, Charles D. Reese, Silvio M. Martino, William T. Carey, Ansel H. Graham

FUNCTIONS OF RAIL HAULAGE SYSTEMS The principal application of rail haulage in under¬ground mining is in the movement of ore or coal pro¬duction from near the mining areas out to a processing plant. In some cases, such production haulage may end at a shaft pocket or slope belt conveyor transfer station, depending on the type of mine and the geo¬logical placement of the mineral being mined. Main-Line Haulage The basic reasons for choosing main-line rail haul¬age are high-tonnage production per shift and long haulage distances. The choice between rail and other systems, such as belt conveyors, is not easy and requires an integrated study of many factors: tonnage, distance, grade, product size, number of producing areas, ore reserve, safety, capital cost, operational and maintenance cost, and the mining system. Projecting a main-line rail haulage system for com¬parison with other systems requires a choice of many variables which are interrelated. The project must start with consideration of the mining system, tonnages per day, distances and grades, and loading and dumping conditions. Rail mine cars of different sizes and types are studied. High tonnage demands consideration of wider track gages. A suggested minimum for metal mines is 914.4 mm (36 in.); for coal mines, 1066.8 mm (42 in.). The choice of the train size should be governed by tonnage per shift, physical factors of loading and dumping, cycle time, and the number of trains that can be operated and safely controlled. After the size of the train is determined, the type, size, and specifi¬cations of the locomotive can be selected. The loco¬motive may be trolley, battery, or diesel; selection guides are discussed elsewhere in this chapter. The advantages of rail main-line haulage are high tonnage capacity, flexibility, reliability, safety, low operational and maintenance cost, and dual usage of the system (supply haulage, for example). The dis¬advantages are high capital investment, high installation cost, the requirement for more maintenance facilities, and restriction to certain ore reserves and mine pro¬jections. Supply Haulage Supply haulage is a basic requirement of any mine, but the type and volume of supplies vary with the type of mineral being mined, the mining methods, and the size of the mine. A mine developed for rail main¬line haulage gains supply haulage through the addition of a secondary track system to extend the main line closer to the producing areas. Additional haulage equipment, generally smaller, is also added. It is neces¬sary to design into the rail system more switch bypass points and train control systems. The track system standards demanded by rail main-line haulage con¬tribute to the reliability and efficiency of the supply haulage function, whereas a supply track system to supplement other haulage systems often suffers from poorer track and installation standards. When belt conveyors are chosen for production haulage, a rail supply haulage system is often installed rather than using rubber-tired vehicles for supply haul¬age. The conditions of the seam or ore body create these items for consideration: haulage grade of more than 6% favors rubber-tired vehicles; longer haulage distances favor rail; difficulty in maintaining a smooth roadway favors rail; and movement beyond the end of a rail system generally requires rubber-tired vehicles, making a single system rather than a dual system attractive. Although operation of a supply haulage system may be concentrated on a single shift of a multiple-shift mining operation, proper planning and supervision are required if the haulage is to be efficient. Often the newer personnel at a mine are assigned to supply haul¬age which may add to the difficulty of obtaining maxi¬mum efficiency. Supply haulage by rail offers the advantages of high reliability, lower maintenance, and high efficiency over long distances. The disadvantages are less flexibility, high installation costs, and rubber-tired vehicles for use beyond the inby end of track often required. Personnel Movement The efficient and comfortable movement of the miner to and from his work area has a significant effect on his daily output. Minimizing man-trip time also becomes more and more important as hourly wages increase. Personnel are generally moved over a rail system by one of three methods: a specialized personnel carrier to which a certain number of men who work in a certain area are assigned; a combination vehicle, used to carry personnel but also capable of being used as a light supply/utility locomotive during a working shift; and man-trip cars towed by a supply locomotive. The first and second methods are used in coal mines where long distances make speed and comfort important. The third method is more common in metal mines. Equipment Movement As mining equipment technology has advanced, many mining machines and their subcomponents have grown in size. This trend requires careful planning for adding specialized carriers to handle the movement of larger equipment. Continuous miners, longwall equipment, raise borers, and ventilation equipment are examples of machines which may be difficult to move and which require special handling. The selection of the weight, speed, and operating characteristics of supply locomotives used to move equipment is important. Larger mines may often justify a special group of loco¬motives for this application.

Jan 1, 1982

There is wide agreement that sustainability has multiple dimensions, but substantially less agreement on the appropriate balance among those dimensions. This occurs because peoples’ vision of a sustainable future, their goals for sustainable development, and their objectives for resource management are reflections of their values, culture, and life experiences. Even assuming agreement about the objectives of sustainable resource management can be reached; there is still a need for policy and decision makers to choose from among alternative courses of action, each of which has its own set of impacts and outcomes. One of several tools used to compare the consequences of alternative actions is the Computable General Equilibrium (CGE) model. This paper uses CGE modeling to demonstrate how policy decisions do not take place in a vacuum and how policy-makers must be aware that there are trade-offs involved in any decision.

Jan 1, 2005

By G. G. le Roux

The purpose of the investigation was to assess various options for increasing the efficiency of the conveyor belt transporting system used at 4 Belt Shaft at Lonmin?s Marikana operations. The reason for the investigation was to prevent the occurrence of a ?bottleneck? at the main ore passes delivering ore to the conveyor belt system, thus creating more fluent operational conditions. The options assessed for increasing the efficiency of the conveyor belt transporting system were: ?Variation of the conveyor speed ?Concurrent tipping from two ore passes on the belt ?Increase the available drive unit power ?Increasing the feed rate from vibratory feeders ?Change in belt width/belt specifications The report shows that the most preferable option to the current situation at 4 Belt, Lonmin, would be the option of concurrent tipping from two ore passes to the belt.

Jan 1, 2008

The propensity of coal spontaneous combustion can be evaluated using various laboratory testing methods such as R70 adiabatic tests, thermal gravimetric analyzer (TGA) and critical self-heating temperature (CSHT) methods. Each of the methods has its unique characteristics and index for classifying the propensity of self-ignition of coals. Accountable relationships among the propensity indices of different methods should be established. Using the newly established R and TGA testing setups in our ventilation laboratory, samples from US coal mines have been collected and tested. 70 The TGA is a relatively new quantitative method. A test can be conducted in a relatively short time. The temperature and weight of the sample can be measured with high precision and then used to determine the activation energy of the coal which is a good indicator for potential of self-heating. Under a specific experimental design, the TGA technique could perform the proximate analysis for determining coal quality parameters (e.g., moisture, volatile matter, fixed carbon and ash content). These parameters can be used in an empirical equation for assessing the potential of coal self-ignition. The trend of activation energies determined using TGA techniques and adiabatic tests are reasonably consistent with each other. As a case study, a comparative analysis of these two methods for determining activation energy was made on coal samples collected from Indiana. The kinetics of the spontaneous combustion mechanism has been explained. If the functional groups of a coal are easier to be activated with a small amount of activation energy at a relatively low temperature, the coal will has a higher potential of self-heating. With the kinetic and coal quality parameters determined by the TGA setup, it has been found that the coal with low potential of self-ignition has low self-heating rate, high activation energy and high minimum self-heating temperature.

Jan 1, 2012

By L. Nahai

THE OUTPUT ff the mineral industry in 1964 in both non-Communist and Communist Europe showed an upward trend for most countries in keeping with the higher index of overall industrial production. Comparable indexes of mineral industry production for non-Communist and Communist Europe (including the U.S.S.R. in Asia) are not available, primarily because the U.S.S.R. does not report details of its industrial production by industrial sectors. However, available data both in the form of index numbers on production published by the United Nations and based on data as reported to the United Nations by the various countries, and actual production data for selected major commodities published by the various European countries reflect generally expanded output by both the extractive sector and the processing sector of the mineral industry in most of the 26 nations of Europe.

Jan 1, 1966

By K. I. Afewu

A modified approach to the implementation of Gy?s sampling theory is discussed and used to design a procedure for sampling crushed gold ore feed to metallurgical process plants. The approach involves calibration of certain ore-specific parameters. These parameters were calibrated for two different crushed gold ores en route to grinding mills in process plants. The model predictions were compared with the predictions from Gy?s original model. Far smaller masses of samples are required by the modified model than by Gy?s model at the same relative variance. The representativeness of the sampling procedure if normally distributed errors are assumed was also calculated. These procedures and evaluations lead to significant savings on cost of handling and preparation of the samples and to a quantifiable degree of confidence in the analytical results for the process plant.

Jan 1, 1998

By H. A. Steane

"This presentation will cover some crushing and grinding considerations in plant design and operation. There is nothing basically new in the subject matter presented here; however, the viewpoint and interpretation may be novel and some of it controversial.The presentation is based on experience gained in the metallurgical design of plants for Placer Development Limited from 1960 to the present. The plants are listed on SLIDE No. 1.These plants have conventional crushing, followed by rod mill and ball mill grinding. I would like to think that autogenous grinding plants are the answer for eliminating secondary crushing plants and reducing the capital and operating costs; however, the autogenous plants seem to have some problems of their own and I have had no experience with them.Fineness of Grinding RequiredFirst let us consider the fineness of grinding required.The fineness of grind will affect the recovery obtained, and the relationship must be determined. If the grinding is too coarse, some of the valuable mineral, either free or locked in middling grains, will not be floated. However, as grinding is probably the greatest single operating cost, we do not want to grind any finer than is justified economically.Finer grinding should not be carried beyond the point where the net smelter return for the increment saved becomes less than the added operating cost. Even this fineness will be beyond the economic limit because of the additional capital cost for the equipment needed to achieve it. These factors will be discussed in more detail later.The Bond work index is the commonest method of expressing a measure of the effort that has to be expended in grinding. It is the kilowatt-hours required to reduce a ton of rock from an infinite size to 80% passing 100 microns. The formula allows a calculation to be made for other size ranges."

Jan 1, 1976

By J. Wennerstrand

Slender steel piles have found an increased use in Sweden in resent years. These piles are often driven throw soft normally consolidated clay to refusal in till or at the rock surface. In contrast to most other piles the bearing capacity with regard to failure of the shaft is in many cases smaller than the bearing capacity with regard to failure of the soil or rock at the lower end of the pile. Imperfections such as deviations from a straight line of the pile centerline and residual stresses in rolled and welded profiles are important for the capacity of the pile shaft. Even thow the second order moment caused by initial deviations in piles is only half of that in a column with the same slenderness, imperfections can and should be accounted for in much the same way as for columns.

Jan 1, 1989

By S. Harteis, D. Dolinar

Many mining operations rely on bulkheads to provide a barrier between impounded water or slurry and active mine workings. However, bulkhead failures could cause and have caused catastrophic flooding that puts the underground workforce at risk. The National Institute for Occupational Safety and Health1 (NIOSH) in collaboration with the Mine Safety and Health Administration (MSHA) is conducting research to evaluate the adequacy of existing design practices for water and slurry bulkheads. A key component of this research effort is documentation of the performance history and design parameters for bulkheads installed in underground coal mines during the last 20+ years. This research is part of a larger effort to develop general design guidelines and procedures for constructing and maintaining bulkheads that will help to ensure their long-term structural integrity, while significantly reducing the inundation risk for miners. This paper presents the results of NIOSH’s research to identify state-of-the-art bulkhead designs, including design criteria, leakage monitoring systems, and emergency warning systems. Underground observations and evaluations of existing bulkheads suggest that the most important design factors influencing their performance are the interface of the bulkhead with the surrounding strata and the potential magnitude of the hydraulic pressure to which they may be subjected. When a bulkhead has failed, leakage has generally been through the surrounding strata or along the bulkhead/strata interface, with the failure potential along the interface increasing with hydraulic head. It was also determined that development of a monitoring program to ensure the long term structural integrity of bulkheads is an important safety consideration. Current monitoring procedures range from weekly visual inspections to constant monitoring via pressure transducers and fluid level indicators. Where possible, these devices are used in conjunction with a computer based mine monitoring system to alert mine personnel when an emergency condition exists at the bulkhead installation.

IN dealing with this subject the intention is not to write a detailed description of the various methods in vogue on the Kalgoorlie field, but rather to give a brief outline of three of the methods which are most universally in force, and from which, ultimately, a few comparisons may be drawn. The three methods which the writer proposes embodying in this article are known as "The Rill," "The Flatback," and "The Shrinkage." For comparative purposes each method will be applied to a block of ore 500 feet in length, and opened up by levels 200 feet apart."The Rin" System. It is necessary in this system, for facilitating mullocking, to make connections between levels by rising and winzing at about every 100 feet or 150 feet along the orebody. After the leading stope is beaten off for the full length ofthe drive, the level is timbered. Single stulls are put in up to a width of fourteen feet. If the span is too great for stulls, set timbers or saddleback stuUs are used. Stulls are lagged over with 4-in gimlet poles, and wooden chutes are built at intervals of 50-feet. Ore passes are cribbed up through the filling directly above each chute, and are constructed of 7-in. round timber or 9 in x 3 in square timber, the passes measuring 4 ft x 4 ft in the clear.Occasionally a double pass may be carried up, one side serving as a travelling way. Stoping is commenced off the winzes, and rills are gradually formed, one on either end of each winze. The rills are generally at an angle of 37° off the horizontal, this allowing for the boring...

Jan 1, 1909

Konkola No 3 shaft experienced severe mining-induced stress and time-related deterioration in ground conditions in an area where an over-cut and bench (OCB) mining method was used to extract a shallow dipping wide orebody. This caused an unacceptable risk for safe mining operations to sustain the required levels of production. After a review of the conditions it was decided to change the mining method to CASCADE to recover ore in the affected area and modify the OCB method and use a modified over-cut and bench method (MOCB) for new mining sections. This paper describes the geotechnical criteria included to derive the mining standards for the MOCB. The geotechnical risks are identified and taken into account to develop stope design parameters, stoping sequences and support requirements. A monitoring programme to assess conditions as mining progresses is also discussed.

Jan 1, 2005

By C. K. Kocsis, S. Hardcastle

Mine ventilation, requires electricity to operate the fans and fossil fuel for heating. Typically, this accounts for approximately 40% of the overall underground energy consumption. This is considerable, and one of the major causes of this expense is that most mine ventilation systems are designed for the "worst-case-scenario" with respect to ventilation demand, which is commonly dictated by diesel equipment usage. Once designed, mines tend to operate their ventilation system at this peak level despite their exceeding day-to-day development and production needs due to lack of appropriate ventilation controls. In underground metal mines, TeleminingTM, which includes the tele-remote operation of mobile development and production equipment from surface, will change the mine design criteria including its support infrastructure such as the ventilation system. With the removal of machine operators to surface control rooms, conventional ventilation systems designed to dilute and remove pollutants can become redundant. Depending on the ultimate need for ventilation, its cost in automated mines with tele-remote equipment operated from surface will be lower than the cost of ventilation in conventional mines. These cost savings, which could be significant, would not only come from the ability to reduce the amount of ventilation in the mine, but also through the capacity to automate its distribution. This paper provides an indication of the new airflow requirements in stopes and dead-end developments for the tele-remote and automated mining processes based on climatic modelling. Furthermore, this paper makes a cost benefit analysis by comparing the ventilation system operating cost of a conventional underground metal mine based on the mobile diesel powered equipment design criteria with the newly redesigned ventilation system operating cost of an automated mine based on the "heat management" design criteria.

Jan 1, 2002

By G B. Ebsworth, J Krokowski de Vickerod

Several drilling programs at Maldon have produced 7790 core samples, which were assayed for gold, with 2825 of the samples assayed for nine other metallic elements. There is a lack of correlation between gold and all other elements. The situation, however, changes dramatically when sorted according to the abundance of arsenic. In the populations where As >200 ppm, there are strong, positive correlations between gold and several elements. There is also a strong positive correlation of Au with the weighted sums and products of all elements with the elemental averages as a weighting factor, and Bi*Sb and Bi*As elemental products. In these populations, there is also a strong positive correlation between the logarithms of Au and logarithms of each element, as well as the elemental sums and products described above. The correlation coefficients increase with higher As cut-off values.All elemental distributions but especially gold show a high-nugget effect and display high skewness and dispersion, low kurtosis, and common outliers. The distributions of their logarithms are several magnitudes less scattered and without outliers. Thus, the logarithmical distribution correlations are best for practical use.Correlation coefficients were T-tested and are statistically significant. The elemental correlations, particularly those of logarithmic values of weighted sums and Bi, could be used as a gold path finding tool at Maldon. After a preliminary assay for gold and arsenic, only samples with As > 200 ppm should be assayed for other elements. Higher values of gold-correlated elements or elemental ratios, in samples with low gold grades, may increase confidence that gold is under reporting due to its highly nuggetty nature.CITATION:Krokowski de Vickerod, J and Ebsworth, G B, 2012. Correlation as a gold path finding tool at Maldon, Victoria, in Proceedings Narrow Vein Mining 2012 , pp 131-144 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 26, 2012

By Lawrence Pelham

Introduction Fluorspar demand or consumption has been taken for granted in recent years, except for those that produce, market, or consume the commodity. Thinking used to be that as the steel and aluminum industry goes, so goes fluorspar. This is much less true today. The industry is becoming more dependent on chemical industries that feed primarily off hydrofluoric acid (HF) production. The US Bureau of Mines (USBM) published an assessment of world fluorspar reserves and resources. It created a foundation of valuable data that can be used to monitor fluorspar availability. Celebrations were held this year commemorating the 100th anniversary of the isolation of fluorine. The US and other countries continue to struggle with the controversy of ozone depletion and the effect of fluorocarbons on the atmosphere. Goals for the government's strategic stockpile are being reevaluated. The market for fluorine chemicals is changing and diversifying, making it a challenge to assess and forecast fluorspar/fluorine demand. US production and distribution Fluorspar Fluorspar shipments from US mines reached an all-time high near the end of World War II in 1944 at 375 kt (413,800 st). Eighty-two mines produced more than 453 t (500 st) each, and 10 mines produced more than 9 kt (10,000 st.) each. Illinois supplied 43% of total shipments. Consumption in 1944 was 372 kt (410,200 st), of which 56% was used by steel producers and 32% by HF producers. This was the last year that US production exceeded US consumption. The post-World War II record for fluorspar shipments occurred in 1951 at 315 kt (347,000 st) 58 mines produced more than 453 t (500 st) and eight mines produced more than 9 kt (10,000 st).

Jan 11, 1986

By W. C. A. van Niekerk, M. H. Fourie

The carcinogenic potential of chromium (VI) has been the subject of numerous environmental and occupational studies, from which many regulatory efforts emanated. Besides the regulation of levels of known toxic substances (e.g. Cr (VI)), the classification of dusts for regulatory purposes has thus far depended on broad definitions of particle size, mainly used to distinguish only between inhalable and respirable particles. In the past, ultrafine particles (less than 0.1 µm diameter) (PM0.1) were often regarded as a non-issue, because it was believed that significant doses would not deposit in the lungs. A growing body of scientific data is indicating that ultrafine particles are more toxic than their coarse counterparts. Increased toxicity can be related to the unique dosimetric aspects of the deposition and disposition of inhaled ultrafine particles. Secondly, ultrafine particles have a larger surface area per given mass. The increased surface area can act as a carrier for co-pollutants, specifically transition metals (including chromium) that could form a coat on the particle surfaces during particle formation. This suggests that the presence of ?inert? ultrafine particles in the environment may enhance the potential for exposure to reactive airborne compounds, including Cr (VI). These factors are particularly important when information on occupational epidemiology is interpreted. It is desirable to consider the particle size distribution, for example when occupational exposure to hexavalent chromium compounds is assessed. Since ultrafine particles are more toxic than coarse particles with the same chemical composition, it could be speculated that cancer slope factors might be related to particle size distribution. This is relevant not only to carcinogenic chromium compounds, but also basically for any particle-associated toxicants. In general, the probability of developing health effects associated with occupational toxicants is likely to be higher when these are associated with the ultrafine particle range.

Jan 1, 2004

Technologies are receiving a lot of 'press' - and rightly so; as they are seen as the means for achieving quantum leaps in productivity and cost reductions. A point of concern for Caterpillar is that we do not believe there is enough understanding of the infrastructure needed to support some of the future technologies, as well there is not enough utilisation of the current systems.

Jan 1, 1997

By M. D. Max, A. H. Johnson

INTRODUCTION Gas hydrates have considerable potential to be a major energy resource owing to their apparent abundance in Arctic regions and along continental margins throughout the world (Kvenvolden, 1988). Yet, the oil and gas industry has been slow to pursue this opportunity. The primary reasons for this are the large conventional reserve base for natural gas worldwide and the perception that gas hydrate will not be a viable resource for another twenty or thirty years. Gas hydrates are thus viewed as a potential resource for the distant future, but not a valid investment for the near term. This paper addresses both global gas markets and gas hydrate technology, and shows that the commercial development of gas hydrate could proceed rapidly enough to constitute a near-term energy resource. GLOBAL GAS MARKETS The proven worldwide reserves of conventional natural gas are truly enormous, totaling over 6,000 TCF according to BP (2004). Moreover, those reserves have nearly doubled during the past two decades and now represent 67 years of consumption at the current production rates of 80 TCF per year (Fig. 1). This reserve increase has occurred even as global production of natural gas has increased by 67% during the same two decades (Fig. 2). While declining conventional gas reserves are a concern in North America, especially in regions where established infrastructure exists to bring gas to consumers, large new discoveries continue to be made elsewhere in the world. Indeed, in many parts of the world (such as the Middle East and parts of West Africa), natural gas has been regarded as an essentially worthless byproduct of oil production or exists in stranded gas fields. Given this situation, pursuing unconventional gas resources that are inherently riskier exploration targets and more expensive to produce would appear to have little rationale from a business standpoint.

Jan 1, 2006

By L. W. Burger

The need to establish and harmonise methodologies to quantify air emissions, atmospheric dispersion and the consequences of air pollutants (especially hexavalent chromium) from the ferrochrome industry became important aspects with the increasing demands for comprehensive impact assessments. The paper includes a synthesis of current knowledge on the following aspects: ? Locally and internationally accepted methodologies for the quantification of chromium emissions, including sampling methods and emission factors; ? A compilation of recognised and accepted air quality sampling and laboratory analysing techniques; and, ? Dispersion modelling and impact assessment tools. Depositing and removing materials from the chromium ore stockpile and wind erosion of exposed stockpiles result in mainly chromite emissions unlikely to contain significant amounts of hexavalent chromium. Pre-treatment processes such as drying, crushing and screening also produces chromium emissions, but mainly as trivalent chromium. The smelting of chromium is the major source of chromium emissions - open furnaces have the highest uncontrolled chromium emissions, followed by semi-closed and closed furnaces, respectively. Based on the current knowledge on emission rates, interim emission factors for particulate, total chromium and hexavalent chromium releases are provided. Ambient sampling and analytical methods to determine environmental exposure of the general public require that hexavalent chromium needs to be quantified in very low concentrations. Suggested sampling and analytical methods for both source characterisation and ambient monitoring are provided. The selection between dispersion models needs to be based on the size of the modelling domain and the complexity of the terrain. The discussion on dispersion models includes the US EPA?s Industrial Source Complex Model (Version 3), the AERMET/AERMOD suite of models and the CALMET/CALPUFF suite of models. Guidelines are also proposed for environmental exposure assessments including cancer and non-cancer health effects.

Jan 1, 2004

South Africa remains the largest global producer of gold, accounting for 11.3 per cent of total world mine production, which amounted to 2520.2 t in 2005. The benefits and disadvantages associated with hedging practices in general, and among South African gold producers is examined. The wide variety and unique nature of risks faced by the mining industries provide a strong case for hedging. Only the principles of hedging, such as selling ounces of gold in the ground at a pre-agreed price for delivery at a future date as a cost-effective mechanism for price risk management, is discussed. Progressive and cumulative increases in the volume of forward sales in the major producing countries are considered by many to be the cause of the declining gold price over the past decade. In spite of negative publicity, forward selling continued as an industry-wide practice until recently (2002) when the global hedge book contracted by 352 t; contraction amounted to 158.6 t in first quarter of 2006. The objectives and benefits of hedging include continuity and stability of cash flows, stability in price achieved, reduction of frictional costs, and creating an effective management tool. The importance of a defined hedging strategy in reducing uncertainty about the hedging objectives is essential to a successful hedging policy. Historical patterns in hedging revenues at an industry level indicate five distinct periods during which the fortunes of hedgers have fluctuated. Finally, since 2001 the rising gold price and reduced price volatility through the consolidation in the global price means that the need to stabilize revenue flows is less urgent. This has meant that hedging is becoming progressively less important for the mining industry at large.

Jan 1, 2007

By J. B. Clark

In 1933, R. paris1 published a phase diagram for the entire Mg-Ca-Zn system. He reported the existence of a ternary intermetallic phase, Mg5Zn5Ca2, which formed a quasi-binary with the magnesium solid solution. Since then, no extensive work has been reported on the system. Recently however, indications of the existence of a second ternary in-termetallic phase in equilibrium with the magnesium solid solution led to a reinvestigation of the solid phase equilibria involving the magnesium solid solution. The 335°C (635°F) isothermal section was chosen for study. Metallography and powder X-ray diffraction were selected as the best mutually corroborating methods for determining the solid phase constitution. Seventy-six alloys were prepared from singly sublimed (99.95 pct) magnesium, electrolytic (99.95 pct) zinc and USP calcium. The analyses of some of the alloys are shown in Fig. 1. The alloys were melted in a graphite crucible under a chloride flu and chill cast in a split graphite mold. Dif-fusion couples (see Fig. 2) were prepared by casting molten Mg-20 wt pct Zn alloy around a cylinder of Mg-30 wt pct Ca alloy. The alloy samples and the diffusion couples were sealed under argon in Pyrex vials and annealed in a furnace capable of controlling within 1/2°C of 335°C. At the conclusion of the anneal, the alloy samples were quenched in ice water and the diffusion couples were air cooled. Procedures standard for magnesium alloys were used for metallographic examination. Powder X-ray diffraction patterns were taken on a G. E. camera with a 14.32 cm dim using nickel-filtered copper radiation. The redetermined 335°C (635°) section of the Mg-Ca-Zn phase diagram is shown in Fig. 1. TWO ternary intermetallic phases, designated w and

Jan 1, 1962