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By Peter Joralemon

Introduction The Sierra Nevada Foothill Gold Belt in California extends from Mariposa north for more than 420 km (200 miles). It includes the Grass Valley and Alleghany districts in the zone of Northern Mines and the famous Mother Lode to the south. The gold belt has produced more than 3.1 kt (100 million oz) of gold since its discovery in 1848. This remarkable production, at the present gold price, would be valued at more than $40 billion. The Law of Gambler's Ruin supposes that averages are based on systematic rather than random variations. It emphasizes the vast differences between finite and infinite probability. By applying this Law to this figure, it can be seen that the belt has produced an average of 15.5 t (500,000 oz) of gold per running mile. Many geologists believe, with some justification, that less than half the gold in this belt has been mined. Sixty per cent of the past production has come from lode deposits, the balance from Tertiary and Recent placer gravels. This paper will concern itself with the developing lode mines along the Sierra Nevada Foothill Gold Belt. As a further and final note regarding placer mining along this belt, Mark Twain suggested that placer miners, having lost their senses, had contributed liberally to the population of California's insane asylums. After a 40-year lull with little or no gold production, the lode mines of this famous belt are showing signs of renewed activity. Two open pit mines, the Sonora (Jamestown) and Carson Hill, are now in operation at substantial scales. Two others, the Goldenbell (Pine Tree-Josephine) and the Royal-Mountain King, are being readied for production on equally impressive scales. All are in the southern part of the Mother Lode. Exploration farther north, at Sutter Creek, Drytown, and Georgetown, is showing encouraging results in developing deeper lode deposits. This new activity has been sponsored to a large extent by junior Canadian mining and exploration companies. Most of the established American mining companies have looked the other way, not participating in this important development. One major reason for domestic disinterest in this opportunity is there is little chance for heap leachable open pit ore in this belt, excluding Carson Hill. And heap leaching has become fashionable with American mining companies today. Also, this great old gold belt must have been worked out, for why else is there so little activity here - Catch 22 logic. Another reason is that if a new deposit of size and grade similar to the mines in the past was found, high costs would preclude a profitable operation. And, finally, California environmentalists would not permit mining, even if a rich deposit were found. These may all be pefectly valid reasons not to explore for gold along this belt. But a look at the history of gold mining in California may suggest that the future for exploration and development of the gold resources in the Sierra Nevada Foothills can be greater than the past. It was during the Depression of the 1930s that this gold belt saw its maximum activity, with more than 3000 mines in operation. Some were father and son operations producing only a few tons of ore a day. Other mines employed hundreds of miners to hoist more than 900 t/d (1000 stpd). Some ores were of marginal grade. Others, like the Sixteen to

Jan 7, 1987

By Gerrit V. R. Goodman, Charles D. Taylor

A technique was developed to determine the fraction of air recirculated from the scrubber exhaust toward the face and the fraction of gas captured by the scrubber. To determine these two values, a series of tests was required. In the first test, gas was released at the face area and concentrations measured at the scrubber exhaust. In the second test, gas was released directly into the scrubber with concentrations again measured in the scrubber exhaust. The simultaneous solution of two independent flow expressions gave values of recirculation and scrubber capture. A series of tests was conducted with various combinations of scrubber flow, curtain flow, setback distance, and machine location. Statistical analyses then were conducted to determine those factors which influenced the levels of scrubber recirculation and scrubber capture in a deep cut. INTRODUCTION As cutting depths increase, the trend is toward the use of blowing face ventilation systems. With these ventilation systems, the potential exists for recirculation between the scrubber exhaust and the intake air stream. One concern is that methane gas, having been removed from the face area, is subsequently recirculated back into the intake airflow. Under these conditions, methane-laden air could be substituted for fresh intake air, leading to increased methane concentrations at the face. There is currently no standard for recirculation between the scrubber and intake airflows. The implied understanding of the Mine Safety and Health Administration (MSHA) is, and has always been, that any recirculation of ventilation air is unacceptable. To minimize recirculation potential, MSHA suggests that operators maintain scrubber and curtain flows within 0.47 m3/sec (1000 cfm) of each other. Given that recirculation may occur in some circumstances, the question arises as to how much recirculation constitutes a hazard for the face workers. Unfortunately, there is currently insufficient information to answer this. Emphasis must instead be placed on means to quantify the amount of recirculation between the scrubber and face area. Methods must also be examined to control this phenomenon. This paper presents a technique for quantifying recirculation and scrubber capture levels in a blowing face ventilation system. Several ventilation parameters are investigated to determine their effects on these levels. DEVELOPMENT OF RECIRCULATION EXPRESSIONS One way to measure airflow quantities at a point near the face would be to measure the gas flow rate Qgas, and gas concentration c at that location. The airflow Qair, would be simply the ratio of these two values, Qair, = Qgas/c. Although gas concentration values, c, can be obtained easily with existing gas monitoring equipment, it is more difficult to obtain gas flow readings, Qgas, at specific locations. To obtain these values of Qgas, another method was needed. Previous work in this area was conducted by Kissell and Bielicki (1975). However, the calculations in their model assumed a constant gas concentration in the region around the continuous miner. In a deep cut, it is questionable that gas concentrations would remain constant or uniform around the mining machine. The technique presented here does not rely on this assumption. It instead assumes that the gas concentration in the scrubber exhaust remains constant. A quantity of gas released at the cutting face either flows into the scrubber inlets or bypasses the inlets and flows around the off-curtain side (in this example, the right side) of the continuous miner (Figure 1). The fraction of the face gas that flows

Jan 1, 1993

By H. Douglas Dahl

Thank you, Don Gentry, for your generous introduction. It is not adequate simply to say that I appreciate very much the vote of the Award Committee putting forth my name as the winner of the 1992 Daniel C. Jackling Award. To look at the list of previous Jackling Award winners is a humbling experience. The Society for Mining, Metallurgy, and Exploration has been blessed with some exceptional members over the years, and I feel privileged indeed to be included among the Jackling Award winners. During the years of my education at Penn State, my field of specialization was mining engineering, with particular emphasis in my master's and Ph.D. work on rock mechanics. But my career has not been quite so tightly focused as that. I have been involved with coal mining in all its variety - its technical aspects, both theory and practice; and its business side, dealing with management and labor, accountants and lawyers, politicians and environmentalists. And I have also been involved with related fields - construction; exploration; barges, railroads, ships and docks; supply houses; the coke business; you name it. Coming from that practical background, I believe it is most appropriate that I talk with you today about some practical matters. But I will not advise you today how to mine coal or protect roof in a hard-rock mine. You fellows know how to do that. The practical matters that I have in mind relate to policy - national policy. I will acknowledge immediately that national policy is often much less than practical, but bear with me as we explore the subject together. I am reminded of a supervisor I had early in my career. He very strongly advised me, back in 1961, to "get involved in the political process," for without objectivity injected into our national policy making, we will become a second-rate economic power. I have selected as the subject of my remarks today, "Coal and the de facto National Energy Policy." There is an energy policy relating to coal, a de facto one, and it's the one we see as a result of regulation, attitudes and culture from our nation's capital. This de facto policy is the one we in the coal business live with every business day. My thesis is very simple - Coal is, in fact, a vital part of our national energy policy and will remain so, as a practical matter, no matter what politicians may do to us, or not do to us. Coal will prevail - that will be the happy ending to my story in which, of course, we are the good guys in the black hard hats. But first, I must flesh out the plot with some other characters and with some threatening activities. There are five main characters in my story: in order of appearance - Perpetual Politics, Rampant Regulation, Selective Science, Electrification and Confident Coal. The time: the present. Perpetual Politics is always a threat to coal. When you cross the Capital Beltway into Washington, D.C., check your wallet and your gold fillings, because you're not likely to run into any friends. On Feb. 20, 1991, just over one year ago, President Bush presented to Congress his National Energy Strategy. It's a thick document of more than 260 pages, with discussions on securing our future energy supplies, increasing energy effi-

Jan 1, 1993

By Arthur F. Eidson

INTRODUCTION If a uranium mill worker is involved in an accident, analysis of excreta for uranium content and[in vivo] counting methods are used to estimate the extent of the possible exposure and to predict the pattern of uranium distribution among tissues (Spitz, [et al]., 1980). Interpretations of urinary excretion data are based on studies of human exposures to single uranium compounds including U02, U308, U03, UF4, U02F2 and UF6 (Alexander, 1974). Refined uranium ore (yellowcake) is a variable mixture of U03-NH3-H20 adducts (ammonium diuranate, ADU) and their thermal conversion product, U308. The variability in composition of yellowcake (Kalkwarf, 1979; Eidson and Mewhinney, 1980; Dennis, Blauer and Kent, 1981) complicates the interpretation of bioassay data from individuals that might have inhaled any single yellowcake material. A test system would be useful to estimate the dissolution behavior of a particular yellowcake aerosol inhaled by a worker and aid in the interpretation of bioassay data collected before and after an accident. A series of experiments was conducted for this purpose. The objectives of these studies were to: 1) develop specific criteria for comparing [in vitro] dissolution systems and 2) use the criteria to evaluate solvents, methods and experimental temperature and pH conditions used in previous studies. Dissolution experiments were conducted using different combinations of solvents, conditions and methods with a common yellowcake sample. Because little information is available from human yellowcake inhalation exposures to evaluate the [in vitro] results, criteria for evaluation of different tests were derived from a model of yellowcake dissolution in lung constructed from a number of cases in which humans were exposed to single uranium compounds known to be in yellowcake (Mercer, 1975). MATERIALS AND METHODS Materials The powder used in the tests was taken from one lot of yellowcake obtained from a commercial uranium mill. The same powder was previously designated as Mill D in Eidson and Mewhinney, 1980 and Mill 2 in Kalkwarf, 1979. The ammonium diuranate content was measured by infrared absorption to be 25% ± 1% (Eidson and Mewhinney, 1980). Solvents The compositions of the two biological fluid simulants used are compared in Table 1. Simulated lung fluid (SLF) (Kalkwarf, 1979; Dennis, Blauer and Kent, 1981) was chosen to simulate the composition [ ] of lung fluid as closely as practical (Moss, 1979). Simulated blood serum ultrafiltrate (SUF) (Eidson and Mewhinney, 1980) was chosen because of previous successful efforts to model human urinary excretion of inhaled plutonium aerosols (Kanapilly Raabe and Boyd, 1974) and the dissolution rate of 241Am02 in Beagle dogs (Mewhinney and Griffith, 1981). Major differences are that simulated lung fluid contains Mg2+ and a higher Ca2+ concentration than SUF and that SUF contains ammonium ion. Both solvents include protein substitutes; acetate is used in SLF, while amino acids are used in SUF. Both solvents include phosphate but the chelating agent diethylenetriaminepentaacetic acid (DTPA) is included in SUF to prevent possible formation of insoluble uranyl phosphate precipitates that would result in an underestimation of solubility. In addition to the above solvents, modified solvent compositions were prepared to investigate the role of specific salts in yellowcake dissolution studies. Solutions of SUF (Table 1) were prepared without DTPA (SUF-DTPA) and without amino acids (SUF-amino acids). No solution excluded both DTPA and amino acids. The SLF solutions were prepared with DTPA (SLF + DTPA) and without phosphate

Jan 1, 1981

By Peter J. King

Demand for both lead and zinc to the year 2000 will be characterised by modest to nil growth in mature economies, with positive 3 to 4 per cent annual growth rates in developing nations. Asia and China will feature significantly in this growth. Overall lead demand is expected to grow about 1 per cent a year, with industry becoming more dependent on the fortunes of the automotive industry as the lead-acid battery will be the only major end use to expand. Zinc will become more dependent on its main use, galvanising, as the automotive and construction industries move to high levels of dependence on corrosion prevention. Strong growth in zinc consumption in S.E. Asia, China and India is likely to feature widespread development of the continuous steel strip mini-galvanising line technology. Thank you Mr. Chairman for inviting me to this Conference. As an Australian I am honoured to be here to offer some thoughts about future trends in metal consumption -- with particular reference to the role of countries in my own neighbourhood, the developing Asian nations, including China. It is customary for a member of the producer fraternity to speak to the topic of future development. This is no doubt due to the fact that our companies are so often engaged in endeavours to determine the future prospects of the metals as essential background to massive investment decisions relating to the development of new mines and smelters. This is sometimes also applied in relation to the more painful decision of closures of aging operations. In this paper I have drawn upon recent studies by AM&S into the future of lead and zinc. While this work has been thorough and skillful, we are nonetheless left with uncertainty on a number of counts. Firstly. predictions in the past in our industry have not been all that credible. Need I remind a U.S. audience for instance of the Paley ' investigation setup by President Truman in 1950 to study U.S. copper requirements for 25 years in the future. Asarco's Simon Strauss gave those of us daring to forecast long term commodity trends a salutary reminder that the Paley Commission with all the expertise at its disposal missed the mark in its copper predictions by a very wide margin indeed. Additionally, it is difficult to take comfort from longer term economic forecasts. Too many important economic issues remain unresolved suggesting considerable caution. These include the threat of growing protectionism; low economic growth rates with stubborn unemployment levels; and levels of government expenditure at near saturation as a percentage of GDP, to name a few. However, on a more optimistic note. we can look to China, where there appears to be momentum in its economic growth; oil, where stable oil prices to the mid 1990's will contribute to inflation control; and the fact that LDC debt, while still a major issue, is declining in importance and priority. In the short time allocated to us. I wish to deal succinctly with the lead and zinc raw materials supply picture, which is one I do not believe will hold too many surprises for this audience. Following the review of raw materials supply, we will turn our attention to the changes in the pattern of demand towards the year 2000. emphasising the Asian position. In this context. I would like to focus on the theme of the changes in OECD consumption levels relative to those of the emerging nations. One area we will be paying special attention to is China.

Jan 1, 1986

By Stanislaw Siewierski

One of the major problems in the exploitation of deep copper ore deposits in the Legnicko-Glogowski Copper Field is ground control and, in particular, roof control of the headings. The depth of copper deposits ranges from 600 to 1200 m. The parameters of ore-bearing rocks and the surrounding rocks show considerable variance of strength properties along the mining thickness (Fig. 1). The roof of the openings exploitation headings consists of dolomite-calciferous rocks. The roof strata consist of layers, the thickness of which vary from several centimeters to 0.6 m (sometimes to 1.0 m). These strata differ in strength and are characterized by igneous and sedimentary discontinuities. On the basis of studies in structure, properties and geological mining conditions of the roof strata in Polish copper mines, the method of roof classification has been developed. (1) This method uses the stability index "L" of the roof, calculated from [ ] The value of "t" ranges from 0, l to 0,8. Based on the value of "L", four classes of the roof (Table 1) corresponding to different degrees of fall hazard have been developed. The mining of copper ore deposit is carried out by a single lift room-and-pillar mining system or a two-stage room-and-pillar system with roof caving. (3, 4). The single lift room-and-pillar mining system is most often used. It consists in working of the deposit of 2.5 to 4.5 m mining thickness, 150-350 m long mining line, consisting of parallel room faces 6 m wide with pillars 5.7 or 16 m wide. As the mining line advances, cross-cuts joining individual rooms are made, thus forming support pillars 3x5, 7x7, 5x16, and 7x16 m (Fig. 2). The two-stage system (Fig. 3) mines the deposit with rooms 6 m wide into blocks 25 x 35 x 25 x 45 m. During the second step these pillars are worked with rooms and cross-cuts into pillars about 3 x 4 to 4 x 5 m (retreat mining). The pillars are crushed just before the fall of the roof. The maximum opening of workings in the two-step system is about 10 m, i.e. one belt entry and one row of bearing pillars. In single lift room-and-pillar systems it is about 35 m., i.e. three belt entries and three rows of bearing pillars. The support of the headings consists of expansion bolts 1.6-1.8 m installed in a bolting pattern 1.1 x 1.5 m. In both systems the fall step is about 10 m, i.e. the width of a belt and a bearing pillar. The Indications of Rock Mass Pressure Individual Dog Heading Individual dog headings are set wide apart from each other in the unmined ore deposit area where there is no interaction of headings. The indication of rock mass pressure are, in such cases, limited to the deformations and damage of side walls. Radiometric studies of rock mass loosening showed that the maximum range of lifts in side-walls immediately after web was 2 m (4,5). In the course of time the cracked side rocks fall off, thus widening the heading. The morphology of side-walls is influenced by the location of the heading in relation to clump layer/tarry. This layer is characterized by a weak strength and "lubricating" properties. The location of the heading in

Jan 1, 1982

By David J. Starbuck

For small precious metals mining to show a comeback a different perspective must be taken than that of larger mining companies. This perspective must be much the same as that of past successful smaller mining enterprises such as when the US had a need for Uranium. Abandoned mines and environmental concerns in older mining districts have discouraged many from entering with the attitudes of why try there if so many unsuccessful mines exist. A small mining venture can be successful there but must have a "think smart'' attitude. There must be a thorough understanding of deposits, utilization of advanced technology and a properly managed mining operation. "THINK SMART" MINING The small mine operator is usually looking for small high grade deposits, consequently, long term success depends on a series of small successful mining operations. The small miner must succeed right-off or credibility and future project funding will suffer. No veteran of even a few years in the mining industry needs to be told that a dollar spent on production today fails to yield as much as it once did. More efficient equipment and techniques have helped soften the problem, but there is little likelihood that a mining company can do much to control inflation, wages, regulations, or grade. That leaves the small miner to "think smart" as one area where operating dollars can still be stretched to produce more than they presently do (Gardner. 1981). In the planning stage of a mining venture, when most options are still open, the main controllable elements are market timing, deposit selection, cutoff grade, mining method, mine design, mining sequence, rate of production build up, process design, project financing structure, project management, and project ownership structure. During planning, optimum criteria can be established for each of these. Once a mine is in production, these options are foreclosed or at least severely constrained. The more thought given to a mining program using all information available the less the risk of encountering disabling surprises. Thus the intent of this paper is to suggest the "think smart" attitude most small mine operations use for long term success. If one looks at a number of successful mining ventures, whether large or small, there is a general logic most have followed to be successful. They have used proven exploration, mining and processing techniques along with economic evaluation and experienced personnel to improve their odds. LOGIC IN DEVELOPING A FEASIBLE VENTURE Most successful small mine ventures achieve positive cash flow as quickly as possible by seeking out those deposits which require the least time and expense on activities leading up to production. To do this they look for deposits that are the most accessible, easily investigated, mined and processed with the cheapest fundamental methods. The general sequence of events starting with district selection are deposit selection, preliminary feasibility, exploration, detailed economic feasibility, mine planning, development and production. Each successive step becomes more costly requiring intermediate decisions along the way. District Selection Small operations should look for known deposits which favor the occurrence of the type deposit sought rather than efforts in unproductive areas. Known deposits in developed areas have some advantages over remote areas: 1. More alternative methods of transportation are available. 2. A larger stable work force is available. 3. The area is usually more pro-mining. 4. Power, water, housing, technical services, skilled services, schools, recreation and shopping are available.

Jan 1, 1987

By J. D. Morgan

Demand for nonfuel minerals depends in large measure on economic activity, which rose in each quarter of 1985. However, despite 1.7 million new housing starts and a 7% increase in motor vehicle production, many domestic metal mining operations experienced continued difficulties as a result of import competition, greater use of newer engineered materials, emphasis on multi-unit housing, and downsizing of motor vehicles. Recognizing the continued difficulties of the domestic metal mining industry, many corporations made major efforts to retrench and rationalize - selling, closing, or spinning off less profitable operations, renegotiating labor contracts, and inviting increased participation by foreign firms. About 15% of US energy was imported in 1985. This included $52 billion worth of crude oil and refined products, adding significantly to the overall merchandise trade deficit of $149 billion. With the exception of a few large-bulk, low-value raw materials, most minerals move relatively freely in world trade. The strong dollar encouraged imports and made exporting more difficult US nonfuel minerals trade deficit totalled $17 billion. The US continued to rely on imports for many important strategic minerals in 1985 (see accompanying Table). Increased pressures for protection included major mineral products such as steel, ferroalloys, and copper. But President Reagan, in his Feb. 1986 "Economic Report to the Congress" stated: "Our international trade policy rests firmly on the foundation of free and open markets." He also stressed the need for fair trade and more open markets internationally. The National Critical Materials Council, created by law in the latter part of 1984, was activated in the latter part of 1985. Its first public hearing was held in January 1986. The Council is to review the nation's materials posture and advise the President on such matters. Compared with 1984, the value of US raw nonfuel mineral output rose 2%, to $24 billion. This supported the production of $244 billion worth of processed materials of mineral origin. Raw steel production - a major indicator of both material use and mineral production - fell 4% to 81 Mt (89 million st). Consequently, domestic iron ore production fell to 49 Mt (55 million st), about 97% pelletized. Iron ore imports fell 7% to 16 Mt (18 million st). Raw steelmaking capacity fell 4% to 117 Mt (129 million st). Pacts with 16 leading steel exporting nations have been negotiated in an effort to reduce import penetration of the domestic market. Domestic producers of bulk ferroalloys of chromium, manganese, and silicon operated at about 40% capacity. The US mined no manganese or chromium. Only a small quantity of ferronickel was produced from domestic laterite mining in Oregon. Molybdenum production increased slightly to 49 kt (54,000 st). Mine shipments of tungsten, another mineral tied to steel production, also fell slightly to 1.1 kt (1200 st) of contained metal. Even the domestic aluminum smelting industry suffered reverses as high stocks and low prices reduced production 16% to 3.5 Mt (3.8 million st). Similarly, imports of bauxite fell 7% and alumina 10%. Production of magnesium metal fell 6% to 136 kt (150,000 st), while titanium metal production remained unchanged at 22 kt (24,000 st). Copper mine production was unchanged at 1.1 Mt (1.2 million st). Copper refinery production fell slightly to 1.1 Mt (1.2 million st). Lead mine production rose 24% to 400 kt (440,000 st). Zinc mine output was down 10%, to 225 kt (250,000 st). Refined lead production increased 28%, to 510 kt (560,000 st). Primary slab zinc production fell 5% to 240 kt (260,000 st).

Jan 5, 1986

By Anthony R. Przybylek, John M. Visnesky

INTRODUCTION Thermal dryers have been utilized for reducing moisture of bituminous coal since the late 1930's. Since that time, numerous changes in the coal industry have motivated advances in thermal drying technology. The purpose of this paper is to discuss the changes that occurred to dryer design and focus on current practices of fine coal drying. The major topics discussed in this paper are: - Early Thermal Dryer Design and Evolution - Current Drying System Design and Associated Components - Current Economics of Fine Coal Thermal Drying - Future Markets and Design Considerations EARLY THERMAL DRYER DESIGN AND EVOLUTION Background The practice of drying fine coal with heated air has been performed for more than one half century. Throughout that period, both Metallurgical and Steam coal producers have used this method for reducing the surf ace moisture content of their fine plant products. During the late 1930's and early 19401s, vibrating bed and flash dryers were used to per- form the drying duties. Closed water circuits and vacuum filtration had not yet been integrated into process circuitry and centrifugal dryers were in their early stages of development. During that same period, changes in mining practices were reshaping the distribution of coal to our processing facilities. Plants originally designed for the purpose of processing coarser fractions of coal were being burdened with additional quantities of fine coal. Those plants, unable to recover fine coal, would reject these solids to plant bleed. The higher incidences of stream pollution that resulted from these events prompted clean stream legislation and the need for new, more efficient fine coal recovery circuits. The coal processing industry responded to these needs by integrating closed water circuits and filtration into circuits. The fine coal filter cakes recovered exhibited poor handling characteristics by themselves and often required further dewatering to improve their characteristics. Flash dryers at that time were being used to dry the coarser products. For those facilities operating flash dryers, the solution to this new problem appeared to be simple. The filter cake was combined with the coarser (-1/4") coal and then introduced to the flash drying column for drying. The filter cakes, however, presented significant problems for the flash drying systems. Heavy lumps of filter cake could not be dispersed and conveyed by the rising hot air and as a result would settle out in the bottom of the column. This material, known as "drop-out", would then accumulate and require further handling. The solution to this problem was not an easy one. Sophisticated mixing equipment was employed to condition the feed to the dryers. As attempts to minimize drop-out with this approach failed, the only recourse was to collect the material with conveying systems and recycle it back to the plant or dispose of it. The additional costs of handling this material reduced the attractiveness of the flash systems and opened the door for the development of a new, more cost effective coal drying system. Due to the nature of the problems experienced by the flash systems, fluid bed systems were considered. Fluid bed dryers were being used in other industries at that time but had no operating history in the coal industry. Dorr-Oliver transferred technology from other industries and built the first fluid bed drying system for fine coal.

Jan 1, 1988

By S. A. Diatchkov

Introduction There is a misconception that mineral-project economic evaluations in the former Soviet Union are either nonexistent or very limited. This is not the case. In fact, there are at least two basic methods that can be utilized to evaluate the vast mineral resources of the Commonwealth of Independent States (CIS). Evaluation procedures are based either on a comparison to known projects, which may rely on limited available data, or on potential-profit calculations, which generally uses more extensive databases. Significant philosophical differences between these systems and those utilized in the Western world include the following: •Under the centralized-government system, the economic evaluation of a mineral project was based on artificially determined mineral prices and costs. Every ten to 12 years, prices for concentrates and metals were established by the Council of Ministers and remained unchanged until the next reconsideration of prices . •The time value of invested money was not considered to be important under these "stable" prices, and inflation was assumed to be nonexistent in the planned economy. •An overestimation of the value of social and economic factors such as keeping the local population employed and providing the country with strategic minerals biased the economic evaluations of projects. Sequence of exploration Geological exploration in the CIS countries typically includes two major phases: prospecting and exploration. Prospecting consists of the following: •Regional geological studies. •General prospecting. •Prospecting. •Prospect evaluation. Regional geological studies are conducted within large territories and utilize satellite images, deep seismic zoning and other state-of-the-art methods. Geological and geophysical maps on a scale of from 1:200000 to 1:1000000 are developed to help identify regional anomalies and mineral prospective areas. The areas of primary interest are then subjected to general prospecting, beginning with geological mapping of the area at 1:50000 (1:25000) scale. The prospecting phase determines mineral potential based on further geological mapping (mostly 1:10000 scale), surface geophysics and geochemistry. Finally, the most attractive targets are subjected to further detailed evaluation (prospect evaluation) to help identify "possible" projects by means of trenching and wide-spaced drilling. The results of the prospect evaluation are analyzed and conclusions regarding the economical potential of the project are drawn. If the initial-prospecting phase results are positive and the prospect is recognized as economical, it then moves into preliminary exploration, which is followed by detailed exploration. Preliminary exploration determines the feasibility of the project on the basis of limited data. For example, in some underground projects, the initial exploration drilling and tunneling is done at two or three times wider spacing than would be required by the assumed mining method. Detailed exploration focuses on preparing the project for mining development, such as completing drilling and tunneling, particularly in the portion of the ore body to be first mined. During the project's mining phase, further exploration is usually conducted along the flanks of the mineralized structures and deeper sections of the ore body. Finally, during exploitation, it is sometimes necessary to delineate the ore body within the mining-block limits. A comparison between the United States (US) sequence of exploration activities and the CIS system is given in the Table 1. There is no "additional" and "exploitation" exploration shown in the US system, but it is common knowledge that all mineral projects are subjected to "additional" and "exploitation" exploration while being mined. An economic evaluation should be performed after the [ ]

Jan 1, 1996

Introduction The US mineral industry in 1983 participated in the recovery from the recent recession. Compared with 1982, the value of raw nonfuel mineral output rose 8% to $21.2 billion. The $218 billion worth of goods processed from this crude mineral output represented a 9% increase over 1982. Even in real terms, these figures reflect growth. Demand for nonfuel mineral materials, of course, depends on the overall state of the economy, which steadily improved each quarter in 1983. Of particular importance to the mineral industry were surges in key mineral-consuming sectors such as construction and automobile manufacturing. New housing starts rose one-third, to 1.7 million units, while automakers turned out about 35% more cars (6.8 million) and 25% more trucks (2.4 million) than in 1982. These are significant increases, even considering recent trends toward smaller, lighter vehicles, and multiple-unit housing. A comeback for the mineral industry, however, is bound to proceed unevenly, particularly at the beginning. Traditionally, the industry lags behind the rest of the economy during a recovery, and some mineral products respond to overall economic improvement more quickly than others. This pattern has been repeated many times and the industry is used to it. Mineral executives know it is best to be patient and cautious during the first stages of an economic upturn. Many experts say that it may take extraordinary toughness, patience, and caution this time around. Steel Improves Raw steel production - a major indicator of both material use and mineral production - rose 14% to 77 Mt (85 million st), stimulating a 7% rise in domestic iron ore production to 39 Mt (38 million It). Still, the iron ore industry operated at less than half capacity, with most major mines remaining closed for part of the year. Iron ore imports dropped 24% to 10 Mt (11 million st), so demand was partly met by withdrawals from stocks. Despite the increase in steel production for 1983, output remained far below levels of just a few years ago. For example, 1983 production was still 29% less than in 1981. Raw steelmaking capacity actually declined about 10% during the year to 122 Mt (135 million st). The industry continued to cite steel imports, especially those from developing nations, as a major cause of domestic capacity loss. As long as steel production remains weak, steelmaking metals suffer. Domestic producers of bulk ferroalloys of chromium, manganese, and silicon operated at about one-third capacity in 1983, with six of the 19 companies involved being closed all year. The US mined no manganese or chromium, and very little nickel, although the country's only nickel mine did resume production in November. Molybdenum production, at 13.6 t (15 st), was only about one-third as much as in 1982. The major molybdenum mines, Climax and Henderson in Colorado, remained closed all year. To meet demand, molybdenum stocks were halved to 19 t (21 st). Mine shipments of tungsten, another mineral tied to steel production, dropped 30% to 1.1 kt (1,200 st) of contained metal. One of the most positive responses to the general economic recovery came from the aluminum smelting industry. By the end of the year, it was operating at about 79% capacity, compared to 59% in January. Nevertheless, the 3.3 Mt (3.64 million st) of primary aluminum produced represented only a slight increase over 1982. Total bauxite imports were about 27% below 1982 levels. Production of magnesium metal rose 11% to 100 kt (110,000 st). But titanium was more typical of other

Jan 5, 1984

By M. E. Rossi

Reconciliation of long-term, block-model reserves and mine production frequently shows significant discrepancies. This reconciliation can be done comparing the long-term, block-model reserves against a grade control model (a model based on blast holes or production data). In this article, this is called an Fl factor for tonnages, grade and ounces. This reconciliation should also be done by comparing the grade control model to the mill feed (or loaded to heaps) tonnages, grades and ounces (F2 factor). The long-term model usually contains blocks with dimensions larger than the size of the selective mining unit (SMU) of the operation. The smoothing effect of an estimator, such as kriging, will generally result in a grade-tonnage curve that does not match the grade-tonnage curve of the SMUs to be recovered. This problem is partly to blame for the reconciliation discrepancies found. In addition, in-pit perfect selection is impossible. Therefore, the grade-tonnage predictions based on blastholes also need to be assessed - considering unplanned dilution and the errors of estimation of the grade control model. This article focuses on using conditional simulations to assess the uncertainty of the predicted grade-tonnage curve in terms of minable reserves and in the framework of long-term mine planning. Quantifying the uncertainty in no uncertain terms is done in conjunction with a popular pit optimizer. It yields several possible minable reserves and several possible minable tonnages and grades, per pit shell or scheduled phase. This is a step further from the usual sensitivity analyses that are performed for deposit evaluation or development planning. Specifically, this approach takes into account the volume-variance relationship, the selectivity of the mining operation and the planned and unplanned dilution. It also carries the local uncertainty implicit in grade estimation into the mine design and mine planning phases. The estimation of recoverable reserves revolves around estimating the block distribution of metal grades from the drill hole information available, typically composites of a fixed length. Ideally, this estimation should accurately predict the ore sent to the processing facilities of an operation. Here are some definitions used in this article: ? Recoverable reserves are those where an attempt has been made to estimate the actual distribution of grades (usually by a block model) that would be found in the ground at the time of mining, relative to the mine equipment size and other operational characteristics of the mine SMU. In particular, it incorporates the effect of internal dilution for each block how the grade distribution changes in variability when comparing a composite grade distribution to the "mining unit" grade distribution. In geostatistical jargon, this is known as "change of support effect," or volume-variance effect. ? Minable reserves are those recoverable reserves defined by a mine plan. In an open-pit scenario, they would incorporate pit optimization in addition to the sequencing and scheduling of production through mine life. ? All other block model estimates from drill hole data that do not explicitly attempt to estimate or measure the volume-variance effect are called in situ. It can be argued that these models may be inappropriate tools for predicting the grade-tonnage curve of the material to be sent for processing (minable reserves). The purpose of estimating recoverable reserves can be thought of as estimating the appropriate amount of dilution that should be incorporated into the model so that it adequately predicts (long-term and on average) the grade-tonnage curve of the ore sent to the processing facilities (mean grades and ton-

Jan 1, 1999

By Spencer R. Titley

The subject of this lecture, a late century perspective on porphyry copper geology, is one of current interest and importance as the search for these ores remains intense. It is presented in the belief that an appreciation of our present understanding of the subject requires it to be viewed historically, thus providing a basis for evaluating appropriate directions for con¬tinued study. At the outset, it is important to note that the technology of mining and extractive metallurgy has developed ore from these low grade orebodies, starting with D. C. Jackling, and has continued to create them as copper grade and price have diminished. But the continued search for them is a geological problem that responds to growing expectations of people and societies; exploration based upon ever increasing knowledge of the habits of the ores has been very successful. This is the story of evolution of geological thought and practice in porphyry ore search across an interval of time beginning with the bulk mining of "low grade copper ores at Bingham and ending with late Century concepts. "How did we get here?" The lecture concerns the ideas of many different people. The ideas and events discussed reflect a bias on North American thought, where the great low grade orebodies were first recognized as such and where many of the basic geological ideas were developed. A very great number of geologists have contributed to this understanding. In no sense is it meant to minimize the contributions to the story that have evolved in the study and mining of ores of the Andes, whose development and expansion has been nearly synchronous with that in North America - nor to minimized comparable contributions made in the late century from studies of the newer discoveries of deposits in other regions, especially the southwestern Pacific regions. In expansion of this topic, I have found that many important concepts are marked by appropriate publication but some are not, or I have not been able to discover documentation. Many notions appear to have become conventional knowledge without publication, mostly being geological viewpoints developed synchronously from similar observation. Thus, apologies are extended in advance to those whose work I have might have missed and to those unknown to me whose unpublished ideas may have seeped into the pool of common knowledge without appropriate recognition. . The analysis here addresses events of two time segments, an older one extending from the developments at Bingham in the early 1900s through the Second World War, ending about 1950; the second period extends from that time to the present. This separation is seen in Figure 1, which traces numbers of publications through the period as extracted from some 700 publications, again with a North American bias, from an annotated bibliography of porphyry copper geology in preparation. During the early period, exploration was "immature" in North America, that is, geologists looked for new districts and sites of the deposits. Outcrops were sought that carried the signatures of former mineralization and regional controls related to structural geology were closely studied. In the later period, North American exploration had "matured" into a study of districts, rather than a search for them. As the studies of these ores intensified during the younger period concommitant with increasing demands for the metal, they did so with a broadened understanding of some habits of occurrence, of hydrothermal alteration, mineral zoning, petrology, and geochemical and geophysical signatures. Outside of North America, search remained "immature" but closely related to the

Jan 1, 1997

By Klimpel. R. R.

Acknowledgments I am honored to be named as a recipient of the Antoine M. Gaudin Award, and I would like to briefly single out four groups who deserve a special thank-you relative to this award. The first group consists of Professor Kawatra, the many international support references that he gathered and the Gaudin Committee itself. I know such things don't come without considerable effort by many individuals, and I'm grateful for their support. The second group consists of the management of The Dow Chemical Company, the managements of various mining companies and, in recent years, the management of G. S. Technologies, all of whom supported me, financially and otherwise, in my grinding, flotation and other particle-processing work. Operations people are the most overworked and under recognized people in the world, whether they be in the mining or chemical industry. Obviously, without extensive industrial support, especially from operating personnel in the mining industry, almost nothing about the work contained in this lecture would have, or could have, happened. The third group consists of a number of coworkers who have played key roles in specifically implementing the new flotation-product portion of this lecture. Of these, Bob Hansen, Don Leonard, Basil Fee, Simon Isherwood, Fernando Cones and Marty Meulendyke deserve special recognition. They have not only been dedicated, but they have been great to work with. Also in this third group are many members of the international academic community, especially those at Penn State. Professor Leonard Austin must be singled out because he took me as a young mathematically oriented chemist and taught me to be an effective process-oriented engineer. Without this experience, I would not have had the over 30-year career as an engineer, which I have loved. In addition, Len planted the size-reduction-research virus in my blood, which I can't seem to get rid of. I also need to mention Professor Frank Aplan, whose obvious love of mineral processing and froth flotation helped to create my interest in flotation. Thus, for those of you who don't like my approach to flotation, it's all Frank's fault. The fourth, and certainly not the least, group to thank is my wife, Charlene, and our two sons, Richard and Steven. No one could have a more supportive family. In every sense of the word, my wife has been my partner from the beginningwe were married at the end of my freshman year in college. She has sacrificed much for me, and I will never really be able to express the depth of my gratitude for Charlene's encouragement, loyalty and support over the past 36 years. General comments on froth flotation The industrial process of froth flotation: Where is it going? I'm here to tell you loud and clear that froth flotation is alive, well, and growing. The inherent characteristics of high mineral throughput per unit of capital, low operating cost per unit of feed, high mineral specificity and high operating flexibility have guaranteed froth flotation an important past and a growing future. As with any evolving technology, there are always trends at the industrial level that are primarily driven by economics. The current trends include: •the movement towards larger and larger equipment, •more alternative equipment designs (e.g., the column cell),

Jan 1, 1996

By G. J. Simandl

Jade is a commercial term encompassing green, white, black or yellow-brown jadeitite and nephrite. Jadeitite consists of Na-rich pyroxene, whereas nephrite consists of prismatic to acicular amphiboles of the tremolite-actinolite series forming bundles that are randomly oriented and interlocked. Nephrite is tougher (harder to break) than jadeitite. Its fracture strength is about 200 MN/m2 whereas that of jadeitite is about 100 MN/m2. On the other hand, jadeitite has a higher hardness (7 compared to 6.5 on the Mohs scale). There are over fifty known nephrite occurrences in British Columbia (B.C.). These occurrences consist of individual blocks, boulder fields, talus blocks and in situ occurrences. Most of the in situ occurrences are lens or cigar shaped. They occur at, or near the contact of ultramafic/mafic rocks (mainly serpentinites) with cherts, and other metasedimentary or igneous felsic rocks of Mississippian to Jurassic oceanic terranes such as Cache Creek and Slide Mountain. These contacts are commonly interpreted as shear/fault related. In general, it is believed that the B.C. nephrite formed by metasomatic exchange between ultramafic and silica-bearing rocks. Impurities in B.C. nephrite are spinel group minerals (chromite, magnetite, picolite), diopside, uvarovite, titanite, chlorite, and talc. Jadeitite is slightly denser than nephrite and good quality material sinks in methyl iodine. There are no known jadeitite occurrences of economic interest in B.C., but there are at least four geographic areas where favorable lithologies, and metamorphic conditions (Pn) required for the formation of jadeitite coincide. These are Bridge River, Pinchi Lake, Dease Lake, and Jennings River. These areas are known to contain small concentrations of jadeite, or are characterized by blueschist or eclogite-grade metamorphic conditions. Rhodonite and pyroxmangite are MnSi03 polymorphs. These minerals are pyroxenoids and are macroscopically and microscopically very similar. For brevity, both are grouped together under the term "rhodonite". Most of the B.C. occurrences are hosted by late Paleozoic chert sequences and are lens shaped and concordant. The common association of rhodonite with iron formations, jasperoids, and/or volcanogenic sulphide deposits in southwestern B.C. permits the delineation of favorable horizons using magnetic or electro-magnetic geophysical methods. Geographic associations between manganese showings and sulphide deposits, and the hydrothermal chemical signature of the rhodonite occurrences from this area, suggest a genetic link between rhodonite and volcanogenic sulphide deposits. The typical rhodonite rocks consist of rhodonite, pyroxmangite, neotocite, black amorphous Mn-oxides, Mn-bearing garnet, ± epidote, ±chlorite, ±rhodochrosite, ± parakutnah-orite, ± Mn-bearing calcite, ± iron oxides and hydroxides. Garnet ± epidote forms yellow patches and layers within the rhodonite. Within some of the rhodonite occurrences, the garnet is altered to chlorite. Rhodonite-bearing rock is commonly cross cut by secondary manganese oxide veins, creating a characteristic web-like texture. In several localities, the contorted bedding planes are readily recognized. Mn-calcite and quartz also crosscut rhodonite. Rhodonite deposits hosted by high grade metamorphic rocks may contain pyroxmangite. Rhodonite is an important gemstone and ornamental stone in B.C., second only to nephrite in dollar value. B.C. nephrite and rhodonite have been used for the past 100 years in local jewelry and ornamental applications. More recently they have been used in table tops, and they have potential as upscale materials in tile-making. There may be a market for tiles made from lower grade materials that were previously extracted and stockpiled. The prices of nephrite, jadeitite, and rhodonite vary widely with quality, but gem and ornamental grade stones are relatively highly priced materials and they trade world-wide.

Jan 1, 2001

By Prescott Greene

AVAILABLE MACHINES Machines discussed in this section include only rubber tire mounted kerf cutting machines of a self¬propelled electrically powered design. Such machines handle bars up to 4.6 m (15 ft) in length, cutting kerfs of from 115 to 191 mm (4.5 to 7.5 in.). The basic design concept of all machines currently in production is very similar. The chassis features dual driving wheels on the cutter-bar end, fixed directly to the main frame. The steering wheels on the outby end of the frame are articulated in some manner to keep all four wheels on the floor over uneven bottom. Machine power for tramming, and all other functions except powering the cutting chain, is furnished from a pump motor that drives the specific pump or pumps required by a given machine. The cutter chain is driven through a mechanical drive train from a separate electric motor having no duty other than driving the chain. The general arrangement of a cutting machine is shown in Fig. 1; two types of machines, with their specifications are shown in Figs. 2 through 5. A hydraulically powered cable normally is included to eliminate the need for manual cable handling and make this a one-man machine. A bug duster, though optional, is needed for most applications. The bug duster is an hydraulically driven scroll conveyor located directly behind the cutter chain drive sprocket. Its function is to move the cuttings away from the sprocket to prevent them from packing up in this area and consuming a lot of machine power to further pulverize them. Available machines vary from models 610 mm (24 in.) high that can cut horizontally only from 203 mm (8 in.) below ground line to 229 mm (9 in.) above, to machines 1143 mm (45 in.) high that can cut in any plane from 178 mm (7 in.) below the ground line to 3.7 m (12 ft) above it. Most currently manufactured machines fall in the 17 200 to 20 400 kg (38,000 to 45,000 lb) weight range, but can be dis¬assembled for underground installation where weight and size are a problem. Maximum teardown results in all pieces having a weight of 5400 kg (12,000 lb) or less and a size allowing passage through a 1.2 x 2.1-m (4x7-ft) opening. Further reduction in either size or weight of the parts becomes a serious problem. Machines currently being manufactured are avail¬able for use with power supplies of 250 v dc; 500 v dc; 440, 550, or 950 v ac (60 Hz). Also available are machines for use on 415, 500, and 950 v 50 Hz. De¬signs to allow operation on other voltages of either 50 or 60 Hz can be made available without any serious problems. As of September 1977, a typical medium-seam ma¬chine without the extra options needed for unusual conditions, but fully equipped to operate, was selling for about $185,000. CAPABILITIES Materials That Can Be Cut The problem of deciding before a machine is put on the job whether it can work effectively is an old one that has never been completely solved and probably

Jan 1, 1982

By William H. Mount

INTRODUCTION The requirements of mine ventilation demand the constant attention of the mine operator. Just as the physical mine structure changes with time, the ventilation requirements of a particular mine change considerably from season to season and even from day to day. No ventilation system can remain adequate indefinitely; and all systems require monitoring and adjustments to achieve proper ventilation. Instead of being given fore- thought and planning, mine-ventilation systems too of- ten are erected hastily, merely to correct an already un- healthy situation. The usual result is an inefficient and costly system that, in some instances, may prove in- effective in providing fresh air to the point of need. The ventilation of mines is a combination of both art and science. The fluid that must be moved is in- visible; its properties change drastically with temperature, barometric pressure, and humidity; and the methods of measurement are not precise. In part, ventilation engineering must rely upon the art of good industrial ventilation experience and sometimes even the trial-and- error method. However, the properties of air can be determined with relatively inexpensive equipment; with some care, measurement accuracies within 5% of the actual values can be obtained. Thus the art of ventilation gradually is being supplanted by improved engineering practices. Although the direction of flow and the volume may not be predicted correctly by airflow equations in every situation, it is better to try to predict the behavior of a given ventilation before installation than to proceed blindly with a system design. When used with the discretion that comes from education and experience, the basic laws of ventilation can provide sub-stantial insight into the design of an air-handling system; and they can prevent wasteful or even dangerous designs from being installed. Since the mine operators who read this text will have varying amounts of exposure to ventilation theory, the basic laws of airflow and of mechanical ventilation are presented in the first section. The emphasis has been placed on the interpretation and use of fan characteristic curves that are supplied by fan manufacturers. Considerable information can be obtained from these fan curves, particularly when they are used with a graphic model of the mine, drift, or duct in which the fan or fans are to be installed. A few minutes spent comparing the various characteristic curves can help the designer obtain the correct fan for the job. However, the under- standing of a few fundamentals is necessary; and this text has been prepared to provide ample and understandable information in a "how-to" style. In subsequent sections of this text, more down-to- earth subjects will be covered, including the problems of air recirculation, relative merits of installing fans in series or in parallel, expected maintenance requirements (vibration, lubrication intervals, etc.), fan noise and some remedies, effect of an inlet bell on the fan per- formance, basic air-measurement techniques and equipment, and similar miscellaneous items. Although this text is not intended to be a comprehensive bible of mine ventilation, it has been prepared as a guide that provides some useful information to individuals responsible for the quality of the underground environment. Every person working underground is en- titled to the safest environment that can be provided within the limitations of existing technology and at a reasonable cost. Adequate ventilation is well within both limits, provided that the system is designed intelligently, in conjunction with the mining plan, and not as an after-

Jan 1, 1982

By Larry S. Diaz

It is felt that prior to the presentation of the technical portion of this paper, some general information would be of interest. The Sar Cheshmeh Copper Mining Company was founded by the Iranian Government to oversee the design and construction of Iran's first copper smelter complex, which will be another step in the industrialization of Iran. Since this is the first such plant in Iran, Sar Cheshmeh wisely chose to enlist the technical assistance needed in regard to the facilities requiring engineering and operating know-how from outside of Iran. To obtain the foregoing, SCCM signed a contract with Anaconda-Iran, Inc., which was formed by The Anaconda Company for this project. The general scope of this contract covers the overseeing of engineering and design by Parsons-Jurden as well as physical plant construction. In addition to this, the same services were given to water supply and tailings disposal systems which were not included as part of the Parsons-Jurden contract. Anaconda-Iran is also assisting SCCM in setting up a training program for Iranian personnel for plant operations, administration, etc., which .will go on for some time after the plant startup, under the guidance of Anaconda-Iran expatriates. SCCM?s other responsibilities, in respect to the development of this industrial effort, are the building of a townsite for approximately 2200 families, the expected manning requirements, having all the facilities necessary to provide a pleasant life in a remote area. It has been estimated that the manpower for operations will be 2050.

Jan 1, 1976

By Aarne Visnapuu

Unmodified pneumatic rock drills may have noise levels of 115 to 130 dB at the operator's position. Experimental modifications have been made to a standard drill to suppress the noise from the three main sources: the air exhaust noise, the noise radiated from the drill, and the resonance noise of the steel drill rods. A prototype model developed at the Bureau of Mines, Rolla Metallurgy Research Center, operates at a noise level below 100 dB. A close-fitting case and muffler unit, consisting of a metallic skeleton filled with viscoelastic absorber and having a durable cover, provided both muffling and noise absorption. Damping alloys were used at strategic points in the drill. Drill steel resonance noise was reduced by a constrained-layer treatment on the outside of the drill rod. The singular and combined effects of the treatments are presented in detail.

Jan 1, 1973

By A. M. Wala

It is common practice in Appalachia, when extending the life of an existing mine, to acquire the adjacent reserves and extend the existing mine into it. Also, often the existing mine is renovated to accommodate the new conditions. The existing mining system and the track and belt haulage systems may be modified while the new property is integrated into the old one. All of these changes create new constraints for the mine ventilation system (M.V.S.) that require proper planning and control procedures. The authors describe and present the study that was carried out for the ventilation system of a newly expanded underground coal mine located in the Appalachian coal region. Discussed first is the process of correcting the ventilation system of the existing mine. Then, some of the stages of controlling the expanded system are discussed. Finally, numerical data for these steps are presented.

Jan 1, 1992