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By Michael Fehler, Hidenshi Kaieda, Leigh House

Hydraulic fracturing has gained in popularity in recent years as a way to determine the orientations and magnitudes of tectonic stresses. By augmenting conventional hydraulic fracturing measurements with detection and mapping of the microearthquakes induced by fracturing, we can supplement and independent1 y confirm information obtained from conventional analysis. Important information obtained from seismic monitoring includes: the state of stress of the rock, orientation and spacing of the major joint sets, and measurements of rock elastic parameters at locations distant from the wellbore. While conventional we1l 1ogging operations can provide information about several of these parameters, the zone of interrogation is usually limited to the immediate proximity of the borehole. The seismic waveforms of the microearthquakes contain a wealth of information about the rock in regions that are otherwise inaccessible for study. By reliably locating the hypocenters of many microearthquakes, we have inferred the joint patterns in the rock. We observed that microearthquake locations do not define a simple, thin, planar distribution, that the fault plane solutions are consistent with shear slippage, and that spectral analysis indicates that the source dimensions and slip along the faults are small. Hence we be1 ieve that the microearthquakes result from slip along preexisting joints, and not from tensile extension at the tip of the fracture. Orientations of the principal stresses can be estimated by using fault plane solutions of the larger microearthquakes. By using a joint earthquake location scheme, and/or calibrations with downhole detonators, rock velocities and heterogeneities thereof can be investigated in rock volumes that are far enough from the borehole to be representative of intrinsic rock properties.

Jan 1, 1986

By A. Bruynesteyn

The results of a 2 phase evaluation program designed to determine the amenability of an oxide ore from the Butte, Montana mine of the Anaconda Company to a sulfuric acid leach is presented. In the phase 1 preliminary evaluation, using roll tests and percolation leaching in small columns, it is shown that the ore is highly extractable at low acid consumption ratios. In the phase 2 confirmatory tests in large columns on l-l.5 tons of material, the amenability of the ore to the leaching process is proven and pregnant solution grade profiles from planned commercial heaps established.

Jan 1, 1978

By T. Ertekin, R. V. Ramani, W. Sung

The emission of methane in the coal mine working faces has two immediate impacts with regard to coal mine planning. Firstly, at the planning stage, the coal mine ventilation system must be designed such that airflows in the faces keep the concentrations of the gas in faces areas to below one percent. If the concentrations were greater than one percent, the production machine has to be shut down resulting in a loss of productive time. In an extreme case, concentration can reach explosibility limits which may lead to ignitions and explosions. Secondly, in excessively gassy coal seams, several alternatives can be applied to the mine layout configuration, and the mining method to be predicted for the extraction of coal, so as to insure a productive and safe operation. In this study a two-phase, two-dimensional finite-difference model was used to predict the methane emission rates into active mine working area under several methane drainage schemes. The numerical model utilizes a recently developed multi-mechanistic flow theory which describes the flow of methane in an heterogeneous and anisotropic coal seam. Transport of methane in the micropore structure of the coal is described by the Fick's law of dif- fusion under the influence of existing concentration gradient, while flow in the more permeable butt and face cleats (macropore spaces) is formulated by Dar- cy's law under the influence of existing pressure gradient. The desorption of the methane is accounted with the aid of a quasi-steady kinetic sorption model. A series of simulation studies were conducted in a systematic manner by sifting through the parameters associated with the coal seam characteristics and the degasification strategies in order to understand those which most significantly affect the methane emission rates in the face areas during mining. Along these lines, the effects of coal seam properties such as thickness, porosity, permeability, sorption characteristics and scheduling of drainage process against the mining operation have been isolated to understand the role of each individual parameter on the methane emission rates.

Jan 1, 1987

By J. Stachulak

This paper deals with theoretical considerations and practical applications of ventilation network calculations, based on mass flow rates and performed by computer simulation in our deepest operating mine - Creighton. The paper also deals with natural ventilation. In the usually accepted air-flow theory, the air is assumed to behave like an incompresssible medium. This introduces considerable error where temperature and pressure changes are great. Therefore, any study of the air flow in deep mines, must take into account compressibility. Basing ventilation calculations on mass flow allows a distinction to be made between air leakage and such quantity changes which occur due to air density changes. The energy which is required to convey one pound of air through a mine can be obtained by adding all head losses hL (ft-lb) of the ventilation paths through which the air flows. However, this total energy will not be correctly expressed in a compressible fluid by summing all pressure losses PL = hL x d (ft-lb/ft3 = lb/ft2) d = density. There is a law of mass conservation, but no law of volume conservation. Natural ventilation is caused by the conversion of heat into mechanical energy and not by the difference in upcast and downcast densities, as often believed. In an existing mine natural ventilation can be determined correctly by plotting a pressure volume diagram and measuring the area ? [ ] vdP = hN, where v = specific volume, or by measuring the sum of all friction losses and work provided by a fan. These procedures are tedious. The natural ventilation head (hN) can be approximated by the equaticn hN =[ ] tdZ. Where T is the average absolute temperature, t air temperature and Z elevation. In Prof. R. Greuer's computer program this determination of hN is automatically done for every mesh of the network.

Jan 1, 1980

By A. Gupta, S. C. Srivastava, K. K. Saroj

Bacterial degasification of methane and conventional methane drainage for generation of power have become accepted practice. Comparative economies of both the techniques have been described. Authors have discussed an alternative technique, involving complete methane biocycle. The first stage of the entire process involves gasification of coal using a consortia of bacteria to generate methane. The generated methane has been proposed to be pumped,as in conventional methane drainage. Technique described is basically similar to underground coal gasification, with the difference that in place of coal combustion, degradation of coal to lower molecular weight compounds by one set of bacteria and then methanogenesis by methanogens, producing methane have been described. Brief account of communities of bacteria that may be employed have been dealt with. Three arguments in favour of proposed technology as an alternative to conventional coal mining are: 1.Total elimination of hazardous operation. 2.Availability of power plants based on methane drainage, assuring comparative economies. 3.Possibility of utilisation of fuel cell technology for power generation giving higher efficiency. All the three phases of complete blocycle viz (1) methane generation from coal (ii) conversion of methane to hydrogen and oxygen (iii) conversion of products so obtained to methane, thereby completing the cycle have been discussed. INTRODUCTION Drainage of methane by conventional pumping process (Boxho,et.al.,1980) not only is helpful in maintaining safe extraction of coal, but with utilisation of extracted gas for power generation, also helps in over all economy of coal mining. This is gradually becoming standard practice for mining of coal in highly gassy mines. An alternative technique (Thakur et. al., 1991) of bacterial degasification of coal seams is also gradually gaining ground. The bye products of this process that is carbondioxide and water, do not contribute towards economy of coal extraction process. However the technique has its own advantage particularly in goaf area for avoiding fire hazard and drainage of methane from coal seams prior to initiation of mining operation. Either of these techniques could be applied in specific situations with success.

Jan 1, 1993

By Dennis P. Bryan

The southwestern United States has an abundance of natural lightweight aggregate (NLA) deposits of volcanic origin that are being used in the construction industry within the region. New Mexico, Arizona, Nevada and California utilize NLA in portland cement concrete and concrete masonry units for weight reduction and related economic reasons. The geological environment of the western United States is amenable to the occurrence of natural lightweight rocks because it contains Tertiary and Quaternary volcanic products that include low-density glass and vesicular material. There are two major classical types of natural lightweight materials of igneous origin; pumice and cinder. These account for the vast majority of volcanic rocks that are presently utilized as NLA. Pumice is an ejecta product generally found as part of tuffaceous units associated with nearby siliceous volcanic centers. Cinder, also an ejecta product, is associated with mafic volcanics and is most often found as a cinder cone. Both are highly vesicular. A third type of natural lightweight material, though also associated with siliceous volcanics, is less vesicular than pumice. It is best described as being a slightly pumiceous rhyolite or rhyolite-tuff breccia that occurs as a siliceous dome or tuft unit that is most likely to be glassy. Uses of these natural lightweight rocks center on the construction industry where they are utilized in lightweight portland cement concrete and lightweight concrete masonry units. The major reason for usage in portland cement concrete is cost saving effected by weight reduction in high-rise construction. The major usage in concrete masonry units, or "block", is for weight reduction that enables a mason to lay more lock.

Jan 1, 1987

By P. A. Roberts

A proposal for a modified in situ leaching method for extracting gold from oxidized gold ores using a non-cyanide lixiviant is described. A non-cyanide lixiviant is suggested because of the obvious concerns posed by injecting cyanide-bearing solutions into the subsurface. Oxidized gold ores were chosen as a focus because earlier research on the use of sodium thiosulfate as a lixiviant under anaerobic conditions indicated that the presence of pyrite led to rapid thiosulfate breakdown. A reconnaissance research program involving ore characterization and hydrometallurgical test work on samples from four Australian ore deposits and preliminary reactive transport modeling studies was carried out. This work showed that lixiviant-oxidant combinations of sodium thiosulfate and ferric EDTA and iodide and iodine are both capable of extracting high percentages of accessible gold from the selected samples in bottle roll tests under anaerobic conditions. The ore characterization and reactive transport studies suggested that both physical and chemical methods of permeability enhancement may be required to lift bulk permeability and the availability of gold for dissolution to sufficiently high levels to obtain adequate gold recoveries. Assuming that such methods prove to be both necessary and economically viable, the mining method would no longer be regarded as simple in situ leaching. Therefore, the term ?in-place leaching? has been adopted for the proposed gold extraction system.

Jan 1, 2010

By W. N. Key, L. C. Acton

Freeport Minerals Co.'s Ertsberg East underground operation typifies the cliche that all the "easy" ore deposits have already been found. Despite its location, in a remote Indonesian rain forest, and the effort and expense required to develop the ore body, the mine was developed on schedule, under budget, and with minimal disruption of the adjacent open-pit mining operation. Development of the Ertsberg East copper, gold, and silver ore body took less than two years after exploration drilling and feasibility studies were completed. In addition, the project came in well under its targeted budget of $101.5 million. Initial block caving and underground ore production took place last year. About 3 kt/d of Ertsberg's total 8.6 kt/d ore production is currently mined underground. This will increase to 5.4 kt/d by year's end. Underground operations will eventually account for all of Ertsberg's production, as the open-pit ore body is depleted. Freeport Minerals is currently studying the possibility of increasing total Ertsberg East production tonnage to 10.5 kt/d by 1983. The equipment, work practices, and development methods at Ertsberg East (Gunung Bijih Timur or "GBT"), while typical of other block caving mines, were modified for rapid mine development and maximum ore recovery. Located in the Irian Jaya province, "Gunung Bijih" is the Indonesian expression for the Dutch "Ertsberg" that in English means "ore mountain." "Timur" means "east" and relates to the fact that the new underground ore body lies about 1 km east of the open pit. The area surrounding GBT is mountainous, with peaks from 4.2-5 km. The underground mine adit and surface plant are located at a 3.6-km elevation. The mine, 4° south of the equator, gets 5.1 m/a of rain, but has an average daily temperature of only 13°C because of the mine's elevation.

Jan 7, 1981

By Peter J. Szabo

The mining executive works in a cyclical industry. His employment is subject to the vagaries of the marketplace. Boom and bust has always been the industry's story. "Big oil" failed in its mining effort. This left emotional and financial scars on many mining families. One must remember that mining executives always play a high-stakes game. Given the cyclical nature of the mining industry, mining executives need to take care of their own financial security. The mining executive that does not plan for his financial security in this cyclical business is playing Russian Roulette with his future. Take a look at those people who have reached their later years with the mining industry. One usually finds three categories of people. Those few in the first group have been fortunate enough to have worked for one company. They have survived the winnowing process of cyclicality, political problems, and takeovers. They have reached normal retirement age. For them, the retirement years are usually comfortable. People in the second group have worked for those same mining companies. But these people have been forced out by early retirement. Or they have been laid off for cyclical, takeover, or political reasons. The third group includes those who have held many positions in the mining industry. Probably not until their later years are they fully vested, if at all, in a pension program. The first small group, then, is the exception. The others, unless they have taken care of their financial security, are doomed to later years of financial problems, stress, and unhappiness. Too late in life, many realize the harsh realities of failure to plan and implement a personal financial program. Mining executives go to some of the world's leading mining schools, to learn how to make money. The curriculum is technically strong. Most mining graduates are familiar with the intricacies of discounted cash flow, present value, and the Hoskold formula. But few of these same mining schools require courses on personal financial planning. Such courses could give mining students a realistic foundation and the mechanisms with which to cope with their cyclical profession. An executive has only three sources of money. He can make money at work. He can put his money to work. Or he can get money from charities, such as government programs. Of these three sources of money, money at work has the best chance of providing financial security in later years. The mining executive must think of the day when he will be too old to work and contribute to the market place. At this point, the quantity and quality of his investments will determine whether his later years will be secure. Along with the Mining Engineering Handbook, two key books should be included in the mining executive's library. These are Venita Van Caspel's The Power of Money Dynamics and Benjamin J. Stein's Financial Passages. These books contain the basic concepts of personal financial planning so critical to an individual's financial survival. Venita points out the simple formula that is the key to financial independence: Time + Money + American Free Enterprise (Rate of Return) = Opportunity to Become Financially Independent Chart 1 shows that this is not a trite formula. The chart shows the dramatic effects of compounding over time. The chart also shows at various yields what a $2000 a year Individual Retirement Account (IRA) contribution will accumulate. Chart 1 (from Money magazine) is based on the future value of an ordinary annuity. It shows that the key to amassing a small fortune by the time a person reaches his 50s is starting to save methodically in his 20s. This, coupled with the highest yields consistent with safety of principal will ensure financial security. There are significant effects from compounding money over time. Every year the decision is put off to save for later years will cost a person dearly by the time he reaches his 50s. For example, consider the decision to put off for one year a $2000 IRA investment at a 12% return. Starting it at age 31 instead of 30 will cost the executive almost $100,000 by age 65. What if the executive waits until age 35 to get started with his $2000 a year investment? He will then lose nearly $400,000 by age 65. Stated another way, the executive costs himself this $400,000 simply because he did not invest $167 a month for five years. $167 a month - that's less than an inexpensive second car payment. This $400,000 pool of money could provide the executive a $40,000 a year income at age 65, assuming a 10% return on invest-

Jan 11, 1985

By Richard P. Graff

Author?s note: Diversity in practice exists in the application of Generally Accepted Accounting Principles (GAAP) for purchase price allocations with respect to the recognition of Value Beyond Proven and Probable Reserves or VBPP, as defined by the Emerging Issues Task Force (EITF) in EITF 04-3: ?Mining Assets: Impairment and Business Combinations?. This diversity appears to arise from differing interpretations of what constitutes sufficient information to value components of VBPP in business combinations. Concerned with this diversity in practice, respected members of the mining community formed a working group which prepared an issues paper that discussed this diversity and provided recommendations for recognition and valuation of VBPP in business combinations. The working group then submitted this paper to the EITF, requesting further guidance be issued on the concept of VBPP. In conjunction with the submission to the EITF, the working group also provided this paper to the United States Securities and Exchange Commission (SEC) for review and comment. The EITF discussed this issue at its May, 2008 agenda committee meeting. However, it declined to add this topic to the EITF agenda for further consideration. Absent additional guidance on VBPP from the EITF, the working group submitted a proposed definition of VBPP to the U.S. Securities and Exchange Commission for its consideration. The following article, which is consistent with the issues paper that was submitted to the EITF and the SEC, discusses the components of and valuation techniques for VBPP. The correspondence with the U.S. Securities and Exchange Commission, which includes the proposed definition of VBPP by the working group, is attached as Appendix 1 to this article.

Jan 1, 2008

By Henry N. McCarl, Gary Waters, William A. Vogely

INTRODUCTION In the past three decades, world trade has grown from less than $100 billion per year in the years following World War 11 to slightly more than $300 billion in 1970 and over $2,000 billion in the early 1980s. Adjusted for inflation, the real growth of world trade in 1972 dollars increased from less than $175 billion per year in the early 1950s to slightly more than $342 billion in 1970 and almost $700 billion per year by the early 1980s. Thus trade as a whole has increased a little more than twofold in real terms in the past decade." Mineral commodities constituted approximately 20% of the value of total world trade in 1970, and significantly more than that proportion in terms of physical volume. By 1980, de- spite rapidly rising petroleum prices, mineral commodities accounted for roughly 18% of the total value of world trade. Crude oil alone ac- counted for about 16%, making it the single most important commodity in international trade. The worldwide recession of 198 1- 1983 brought a dramatic decline in both volume and price of crude oil, shrinking the impact of mineral commodities to roughly 10% of the value of total exports, worldwide, in 1982- 1983. The decline of mineral commodities as a pro- portion of the value of world trade in the decade 1970- 1980 is due largely to the increasing importance of manufactured goods and consumer products such as automobiles and electronic items. During this period, mineral commodity trade expanded roughly sixfold in terms of current dollars to an estimated $350 to 360 billion in 1980. This was equivalent to doubling the real value of mineral commodities in international trade in the 1970- 1980 period in constant dollars. The U.S. Bureau of Mines has expanded the traditional definition of mineral commodities to include semifinished products such as iron and steel (in addition to iron ore), refined metals such as copper, lead, zinc, aluminum, and tin, concentrates, scrap, and petroleum products. Since large quantities of essentially mineral derivatives travel in this form, it is reasonable to use these expanded definitions in stating the role of minerals in world trade. Thus minerals in world trade comprise 30 to 33% of total export trade in all commodities during the period 1976- 1980 (see Table 1.3.1). In addition, the U.S. Bureau of Mines estimates the value of all mineral commodities consumed in the world (at the national level) and compares this value to total worldwide commodity consumption. By this system, mineral commodities range from 30 to 40% of all commodities consumed each year in all countries during the period 1976- 1980 (see Table 1.3.2). The growth in export trade in mineral commodities by commodity group is shown for the period 1976-1980 in Table 1.3.3. These data parallel the swings in worldwide economic activity during the same time period. By any of the above measures, minerals, especially the fossil fuels, constitute a significant proportion of both total economic activity and

Jan 1, 1985

By Leo Misagi

THE ROLE OF MANAGEMENT IN PLANT SAFETY Safe Production In plant management, safe production is an ideal. It becomes reality when: 1. Plant employees work in healthful and safe conditions. 2. Plant safety is in harmony with productivity. 3. Worker absenteeism and turnover are low-the work force is stable. Plant management must apply its talents, skills, and resources to build and maintain the foundations of safe production on a system of policies, procedures, and attitudes. It must do so by marshaling all managerial functions - staffing, direction, control, innovation, and representation-toward that goal. Ideally, a plant safety program is proactive. Management should plan and implement safety activities that are aimed at prevention. Past accidents-accomplished facts-should not be the sole guide of future actions. In other words, positive preventive measures, not the fewness of accidents are the criteria for safety production. An ideal safety program is one which is cost effective. Management should view plant safety as part of a broader program-loss control. Loss control includes prevention of fires and other accidents, unhealthful conditions, and occurrences that cause bodily injury, illness, property loss, or damage. Management evaluates the employees' productivity and safety. An effective training program aimed at loss control is the hallmark of safe production. Ideally, both labor and management should be well versed in safety. They believe in the premise that work-related injury and illness, as well as property loss and damage from fire and other accidents, are manageable and preventable. Such belief prompts them to learn more about the work process and potential hazards associated with each job. They also learn how to expect and prevent fues and other accidents, and respond to emergencies at the plant. Plant management determines which individuals or groups of employees are accountable for the following functions:

Jan 1, 1991

By Kevin Sharman, Peri Mehling

A detailed assessment of acid rock drainage (ARD) potential was conducted at the Quintette Coal Mine site in northeastern British Columbia The site is unusual in that the coal seams contain low sulphur levels as compared to eastern coals (typically less than 0.5%), and the non-coal wastes generally do not to have the potential to generate acid on the basis of static ABA tests. The presence of small iron-stained seeps on the pit walls indicate localized sulphide oxidation, while the overall mine site drainage has remained of excellent quality with extremely low levels of metals after approximately 15 years of mining. Thus the site provides an excellent comparison to more problematic coal mine sites. For the Quintette assessment, static acid-base accounting (ABA) analyses of individual samples were compared to criteria for potentially acid generating materials developed for the U.S. Appalachian coal, as well as criteria developed in British Columbia for hard rock mines. Geologic units were characterized by the percentage of samples, or weighted thickness, that was deemed potentially acid generating (PAG). Links were then made to actual seepage and water quality emanating from geologic units around the site. Seeps at the site demonstrate a large range in pH (2.32 to 8.18). The few acidic seeps (pH < 4.0) were found to contain elevated levels of sulphates, dissolved iron, manganese and aluminum, consistent with the metals typically of concern in Appalachian coal mines. Other heavy metals that were detected in acidic seeps at Quintette included copper, zinc, nickel and cadmium. The use of simple ABA tests, and comparison of individual sample results with the Appalachian and British Columbia criteria, was found to be useful in predicting the probable sources of the small localized acidic seeps.

Jan 1, 2000

By Nora M. Hutt, Kevin A. Morin

Geochemical static tests are a basic component of geochemical assessments and predictions for acidic drainage, and include acid-base accounting and total-metal contents. A key database for such information is the International Static Database (ISD), which currently contains more than 45,000 static-test analyses from over 200 minesites. Such large databases are useful in identifying general trends and, just as important, the percentage of exceptions. General trends and significant exceptions can be seen in the ISD by comparing one parameter against another. This paper shows that bulk neutralization potential (bulk NP) obtained from acid titration is often a good indicator of neutralizing capacity above values of 10 t CaCO3/1000 t. Other types of calculated neutralizing capacity are then compared to bulk NP to determine if they, in fact, reliably represent actual neutralizing capacity and if they can substitute for the more time-consuming and expensive bulk NP. For example, bulk NP correlates better with Carbonate NP (from carbonate analyses) and Total-carbon NP (from total-carbon analyses) than with (Ca+Mg) CaNP (from solid-phase analyses of calcium and magnesium). This means that, in general, (Ca+Mg) CaNP is a less accurate substitute for bulk NP than the other types, and thus requires intensive site-specific justification if a substitution is attempted. This also means that measurement of the frequently less-expensive parameters of total carbon and carbonate can often substitute, within site-specific error bars, for the bulk-NP titration. However, the ISD also shows that non-carbonate minerals make the primary contribution to neutralization in up to 20% of samples, so Carbonate CaNP and Total-carbon CaNP do not always reveal the full neutralizing capacity of a sample.

Jan 1, 2000

By O. Guyot, A. Broussaud, J. Björkman, P. Söderman, P. O. Samskog

Model-Based Expert Control systems (MBEC) are on-line supervisory dynamic optimization systems which combine the ability of a phenomenological model to generate a large amount of information about the state of the process and predict its evolution, with the flexibility of a rules-based approach to take plant constraints into account and apply a complex control strategy. For its new concentrator started in 1995 in Kiruna (Sweden), LKAB has implemented a progressive grinding flowsheet with closed-circuit autogenous grinding followed by closed-circuit pebble grinding with ore pebbles. The standard for instrumentation in this new plant is very high. The control hardware consists of a Distributed Control System (DCS) through which the operators interact with the plant, supplemented by a site-level Management Information System (MIS). A state-of-the-art MBEC system was imbedded in the MIS by the authors before plant start-up. It consists of several modeling and expertise software modules: a phenomenological process model and a Kalman filter estimate continuously the state of the process, an optimizer calculates an optimum feedrate, while rules-based modules establish production targets and key process set-points. The decision to install the MBEC system as an integral part of the new plant before start-up illustrates the current maturity of the technology, i.e., enough know-how was accumulated through former projects to design and develop the system to a large extent without accessing the plant. The feasibility and the power of dynamic optimization systems were previously demonstrated in Kiruna pre-existing concentrator by the authors. First, a MBEC system was installed on the three existing pebble mill circuits, providing both an increase in quality (stabilised size distribution) and a 10% increase in capacity. An on-line expert control system with limited modeling capabilities was also installed in the existing flotation plant, with the aim of stabilizing phosphorous content in the concentrate and optimizing collector additions. Both objectives have been achieved.

Jan 1, 1995

By K. S. Venkataraman

The physical properties of clay-water systems depend on the complicated system of forces between the clay particles themselves, and between the clay particles and the ions in the liquid phase. The kind and distribution of ions in, on, and between the clay particles and the size and the shape of the particles are the basic factors determining the macroscopic behavior of clay-water systems. Understanding the system requires a knowledge of the nature of the clay particles, their size, structure, composition, and surface properties, and of the manner in which they interact with ions [and molecules] in the surrounding liquid [or other medium]. The validity of Professor Brindley&apos;s words (Brindley, 1958), written three decades ago in the context of making pottery, whitewares, and electrical porcelains, transcends time, and the basic message is perhaps all the more important in the considerably expanded use of ceramics for structural, thermal, tribological, electronic, and other applications. Silicon carbide, silicon nitride, and sialons have been studied in the last two decades for high- temperature structural and tribological applications, particularly for using in internal combustion engines. Titanates, zirconates and niobates of barium, strontium and lead, have high dielectric constants, and are extensively used in the formulations for making capacitors. Hexagonal ferrites (molecular formula MO.6Fe2O3) are in use for making permanent magnets for fabricating miniature motors, and for assembling loud speakers, particle accelerators etc. Cubic ferrites such as magnesium-zinc ferrite and nickel-zinc ferrite are used as transformer cores, and for other high-frequency applications. In this context, Richerson&apos;s recent book (Richerson, 1984) on the general scope of traditional and technical ceramics is a good starting point for an overview of contemporary ceramics technology. Glasses are a whole class of amorphous materials used widely as sintering aids, and for making glass-bonded ceramics and glass-ceramic composites. Composites are yet another burgeoning field where two or more particulate components are used for improving the performance of ceramics. For all these applications, the inorganic starting materials are almost always submicron and near-micron powders. Understanding the powders&apos; physicochemical properties, and their surface chemical interactions with the surrounding liquid/gaseous medium is-necessary for making reliable ceramic parts at competitive prices. Even though ceramics science and engineering has attained its separate identity in universities and the industry, ceramists themselves would concede that ceramics science is a cross-disciplinary field, having incorporated and assimilated within itself many principles from several apparently disjointed disciplines. Principles of material science, graduate-level physics and chemistry, polymer science, surface and colloid chemistry, transport phenomena, particle technology, unit operations commonly used in chemical engineering and mineral processing, and statistics and applied mathematics are integral part of any ceramics curriculum in universities. Added to this is the fact that all bench-scale successes in making ceramic parts are to be scaled-up for larger throughput operations. Understanding and applying process engineering principles of comminution, classification, drying, calcination, etc. then becomes essential. CERAMIC FORMING: Despite the diversity of the materials and processes, conceptually, the steps involved in making ceramic parts have remained the same over several decades: The different components for making the pan (usually one or more powders plus other forming and sintering additives) are proportioned and mixed thoroughly, and the well-mixed formulations are consolidated into desirable shapes known as "green bodies." Usually binders such as wax, clay, organic polymers and surfactants, whether dispersed or dissolved in a suitable liquid are used during mixing the batch for giving strength for the green bodies. In the dried green state, the inorganic powders typically occupy only 55 to 60% of the bulk volume of the body, depending on the particle size distributions of the powders and the forming history, with mostly inter- particle voids accounting for the rest of the void volume. SINTERING: The formed bodies are then fired in high- temperatures kilns/furnaces during which the parts are exposed to a predetermined temperature profile, and "soaked" for a certain duration at the final high temperatures, typically between 1200 K and 1900 K, and then cooled to room temperature. The gaseous atmosphere in the furnace is controlled (oxidizing, reducing, or inert) when necessary. During the initial stages of firing, volatile liquids evaporate, and during the intermediate temperatures between 400 and 600 K, the the organic polymeric additives pyrolize and oxidize into water vapor, CO, C02, and other gases. At still high temperature, the glasses, when present, soften, and simultaneously, the ceramic particles rearrange into a network of grains with definite grain boundaries so as to reduce the total interfacial free

Jan 1, 1990

By R. J. Pindred

Discussion by R.J. Pindred In his paper, Davis presents an overview of risk. He also introduces the Capital Asset Processing Model (CAPM) as a foundation for selecting the appropriate discount rate for a mining project. While applying portfolio theory is more defensible than the ad hoc adjustment of discount rates, the CAPM is not a panacea. CAPM shortcomings [The CAPM, as Davis stated, is expressed in the equation: ri=rf+pi4) where ri is the project discount rate rf is the risk free interest rate (3i is the project beta, and 0 is the market risk premium (rm - rf)] Application of the CAPM is more difficult than Davis indicates. Valuation is prospective, while the CAPM parameters are historical. Beta is determined from a regression analysis of historical data, while the beta needed for valuation is the expected beta. Betas are known to be unstable and the regressions that generate them often have low explanatory power. The difficulty of estimating a "project" beta must also be considered. Thus, the beta that is used in the CAPM will be based on the analyst&apos;s judgment. Like Cavender&apos;s discount rate, this judgment can lead to different project NPVs. Subjectivity in valuation cannot be avoided by a mechanical application of the CAPM. The risk-free rate, which Davis identifies as a short-term real rate of 4%, is also subject to scrutiny. A mining project is not a short-term investment and no single risk-free rate is appropriate for all of the cash flows. The hypothetical mine discussed in Cavender&apos;s paper is a six-year project. One might argue for the application of a risk-free rate from the Treasury yield curve at the duration of the project (in a bond-duration sense). This, too, is inappropriate. The risk-free rate should be matched to the timing of the cash flow. These rates can be determined by calculating the implied forward rates from the yield curve using a procedure known as "bootstrapping." It is likely that each of the project&apos;s cash flows would be discounted at a different rate. Commodity prices Davis criticizes the "ad hoc adjustment to the discount rate." Yet, in his discussion of the value of stochastic simulation, he suggests that the gold price be modeled as a "random walk, with or without a trend." This is essentially an arbitrary modeling of price risk. Consider that a liquid market in gold futures exists. The futures&apos; price curve, which is closely related to the market&apos;s estimate of future spot gold prices, should be used to provide inputs to the model. This is especially true of a relatively short six-year project. Alternatively, as Davis correctly points out, a risk-averse investor can sell the commodity short to hedge price risk. Is it any more correct, in the portfolio sense, to account for price risk at all ?? References Cavender, B., 1992, "Determination of the optimum lifetime of a mining project using discounted cash flow and option pricing techniques," Mining Engineering, Vol. 44, No. 10, pp.1262-1268 Fabozzi, F.J., 1993, Bond Markets, Analysis and Strategies, Second Edition, Prentice Hall, Inc. Higgins, R.C., 1992, Analysis for Financial Management, Third Edition, Richard D. Irwin, Inc. Solnik, B., 1991, International Investments, Second Edition, Addison Wesley Reply by G.A. Davis Pindred discusses two issues related to my paper, the shortcomings of the Capital Asset Pricing Model (CAPM) and which commodity price values to use in the valuation exercise. Even though these topics are not directly related to the use or misuse of Monte Carlo simulation, they are important points to take into consideration in valuation exercises. Since I do not appear to have addressed these issues satisfactorily in my original paper, I will comment on each here. Pindred agrees with me that applying portfolio theory, and specifically the CAPM, to the selection of project discount rates is more defensible than ad hoc methods. But he then points out that the application of the CAPM to project valuation is more difficult that I indicate. It is true that the CAPM is a difficult tool for project valuation in general,. But the application of the CAPM to mining projects is one of the easiest I can think of. The biggest problem with using the CAPM for project valuation is coming up with an expected project beta. I suggest a project beta for gold projects of 0.45. The "true" value might be 0.35, 0.55 or whatever. Pindred correctly notes that the selection of the appropriate project beta is based

Jan 1, 1996

By P. R. ZETTWOOG

HISTORICAL BACKGROUND A program in France to develop personal [a] dosimeters has been initiated 1974. The patent on which is based the present device was obtained in 1972 * . From 1972 to 1974, the possibilities of applying certain ionograph track detectors to the spectrodosimetry of radon daughters was explored. The first prototype were produced in 1974. It took four years (from 1974 to 1978) to produce an autonomous dosimeter whose components has a sufficient life span, especially for the turbine motor unit. Qualification in the laboratory was obtained in 1977. In 1978 it was obtained in the mine for technology (autonomy of 12 hours and a life span of more than one year) and in 1980 for monitoring. 300 dosimeters have been tested in underground mines all together. Indispensable peripheral equipment were also developed from 1976 to 1980 : calibration devices, equipment to prepare and develop the films, read out systems. The concept of an "Integrated System of Individual Dosimetry" (ISID) based on a personal [a] dosimeter measuring exposure to radon daughters, thoron daughters, ore dust and external irradiation doses was proposed at the end of 1980. Since January 1st 1981, ISID is used on a routine basis in some french mines, situated in remote area, and appears to be very competitive with the ambiant dosimetry. The latest version of the dosimeter is produced in mass series since June 1981 and should equip all french mines in 1982. DESCRIPTION OF THE INSTRUMENTATION DEVELOPMENT OF THE DOSIMETER MEASURING HEAD The measuring head is based on the use of ionographic film to detect a tracks. In fact, the measuring head is a spectrodosimeter which measures separately over the period of exposure: - the potential [a] energy inhaled due to the decay of Po 218, Po 214, and Po 212 ; - the number of Rn 222 atoms inhaled ; - the inhaled total [a] activity of the five long-lived emitters present in the ore dust. The contribution to the total inhalable potential [a] energy of these various radionuclidesin a typical underground mine is studied in Appendix I. The measuring head described in detail in Appendix II, is able to satisfy all the implications made in the ICRP recommendations. Appendix III deals with the use of this measuring head in the cases where the equilibrium factor is lower than 0.1. This situation occur in open-pit mines where account must be taken of the Rn 222 contribution, which is no longer negligible in relation to that of its daughters. CURRENT DOSIMETER PERFORMANCES Table I shows the characteristics of the latest dosimeter. Appendix IV should be consulted concerning qualification of the dosimeter in the laboratory and in mines, technological development which finally produced the [a] dosimeter and its peripheral equipment, and technical presentation of the ISID (Integrated System of Individual Dosimetry) based on the concept of a multirisk personal dosimeter. Data on the installation and operating costs of such a dosimeter, which would seem to be competitive, are also given in this Appendix. ADVANTAGES OF PERSONAL DOSIMETRY AS COMPARED TO AREA MONITORING The results of the first eight months of experiments carried out under real conditions in an underground mine site are given in detail in reference 8. Area monitoring : The monthly exposure per worker to inhaled Rn 222 was determined from the knowledge of time spent in various areas of the mine and for the different mining operations, as well as from numerous and systematic sampling of the Rn 222 concentration in all work places. Personal dosimetry : The exposure to potential energy from radon daughters was measured by an [a] dosimeter developed by the CEA and worn by each of the 160 miners during eight months. In this way 160 x 8 pairs of monthly individual exposure values have been obtained which can be statistically studied. This test was decisive for us because it proved that the [a] dosimeter was technically sound (very few defects over one year for 160 dosimeters) and especially that personal monitoring devices were superior to area monitoring devices. The following conclusions can be drawn 1. The exposure distribution obtained by personal dosimetry is log-normal. This is true for the results on the whole as well as for groups of results relating to certain explanatory variables. See fig. la, 1b, 1c. 2. The exposure distribution obtained by area monitoring does not correspond to any type of distribution. If the results of personal monitoring are taken as a reference, area monitoring tends to underestimate the high exposures and overestimate the low exposures.See fig. 2. 3. [a]-energy exposures are underestimated when calculated from radon exposures and the equilibrium factor found in the considered mines. This is due to episodes or to zones of high radon concentrations not registered

Jan 1, 1981

By Victor F. Hollister

Skam deposits considered as porphyry copper deposits have been classified into three basic prospecting models: (I) the skam deposit associated with a mineralized intrusive (Pima-Mission), (2) the deposit associated with a barren igneous stock (Ertsberg), and (3) mineralized rock that seems to have no clear association with apluton (Phoenix). Pima-Mission, Christmas, and many other copper-bearing skarns developed in carbonate sequences near mineralized intrusives form an important part of the deposit. Historically, copper derived from the mineralized intrusion accounts for less than half of the metal produced by most porphyry mines, as most mineralization occurs in the intruded rocks. Skam mineralization is an important part of wall-rock metal deposition, because carbonate sequences are preferentially replaced by sulfide and skarn minerals. Exploration for skam mineralization is an important phase of porphyry copper exploration. Impure carbonate beds can become better host rocks for skam deposits than pure carbonates, and these in particular should be included in porphyry-exploration programs if the impure strata are cut by the stock. Skam deposits can also form near stocks that are barren of sulfide. In this model, the two best ore guides are strata favorable for mineralization (usually impure carbonate beds) and the intrusive contact. Faults can also act as important ore controls. Some skarn deposits form that have no clear genetic connection with a stock, although intrusive rocks are known in the district. In such cases, the best ore guides are strata and district-wide faulting. Dikes and sills are also common and mineralization can locally prefer their contacts even though the igneous rocks are barren. Skam minerals are influenced by the chemical composition of the carbonate. Forsterite, brucite, tremolite, and talc are common in skams developed in dolomitic beds. Skams derived from calcic beds often contain wollastonite, calcite, stilbite, or idocrase. Diopside, however, is a common accompaniment of copper sulfide in skam regardless of the original composition of the carbonates. Tremolite and possibly other amphiboles and chlorite can occur significantly in the zinc-lead portion of skams. Gamets occur in magnetite, copper, and zinc-lead zones of skams, but are not as diagnostic as diopside for the copper zone or tremolite for the zinc-lead zone. A general desciption of skam-deposit mineralogy can be used to construct prospecting models of porphyry copper skams. An abstract of such a description follows.

Jan 1, 1991

In 1989, United States mine production of barite decreased 13% while world mine production decreased by 8.6%, according to the US Bureau of Mines. Consumption, sold or used by grinding plants, decreased by 21.2%. And imports decreased by 12.8%. The value of domestically produced barite, fob mine, increased 17.4%, according to the Bureau. The declared value cif US port of all imported crude barite during 1989 increased from $38.10/t ($34.56 per st) in 1988 to $40.27/t($36.54 per st). Nevada continued to be the leading producer of barite with 77% of the total, followed by Georgia and Missouri. The Bureau of Mines estimates 75% of the domestic mine production was used as a weighting agent in drilling fluids. The remaining 25% was used in barium chemicals, glass, or as a filler. Most of the production from Missouri, Georgia, and Tennessee was used in the nonoilfield sector. Of the total consumption used by grinding plants and chemical manufacturers, it is estimated that 80% was used in drilling fluids. The other 20% was used in the non-oilfield sector, including barium chemicals and fillers in paper, plastics, and glass. The tonnage consumed in the drilling fluids market decreased due to the lower average rig count during 1989. The average number of rotary rigs operating was 869, 7.2% less than the average of 936 in 1988. Crude barite imports were down 12.8% from 1988. Imports accounted for 73.8% of the barite consumed (apparent) in the US. China continued to be the leading exporter of crude barite ore to the United States with 63% of total crude ore imports. India followed with 27.4%, Peru 4.3%, Chile 2.9%, and Canada and Mexico, each with 1.1 %. About 97% of the imported barite ore was used in the oilfield sector.

Jan 1, 1990