Search Documents

By Tim Neil, O&apos

Acid rain legislation, the new tax package, excess coal capacity, the effects of low oil prices, how to increase coal exports: These were among the items discussed at the May 4-7, American Mining Congress coal convention in Pittsburgh. Some 2000 people attended the convention, which also offered 15 technical sessions. As always, the state of the domestic coal industry might be characterized as "long-term promise, short-term problems." And one of these problems is acid rain. Acid rain The proposed acid rain legislation in Congress could be the most costly piece of environmental legislation ever written. In its present form, the measure could cost the nation up to $110 billion over the next 15 years. Rep. Henry Waxman's (D-CA) bill, HR 4567, would mandate large reductions in sulfur dioxide emissions from coal-fired power plants. The bill has more than 150 Republican and Democratic cosponsors. Ed Addison is president of the Southern Co., one of the nation's largest utilities and users of domestic coal. Addison noted that America's electric utility industry buys and uses nearly 85% of the coal consumed in this country. He said Waxman's bill would drive up prices of low-sulfur coal, raise electric rates, and force miners out of work in high-sulfur coal regions. In repeating a standard coal industry response, Addison said the Clean Air Act is doing the job. In recent years, while coal use has gone up, S02 emissions have gone down. Current air pollution standards are producing cleaner air, he said. Despite concern over HR 4567, the bill's future is uncertain. Several coal industry executives and analysts predict the bill will die under weight of opposition from coal, utility, and steel interests. But the acid rain issue is gaining momentum. Future legislation of some kind is likely. Meanwhile, research continues to develop clean coal technology to deal with the S02 problem. Commercialization of these front-end technologies currently lags public sentiment for acid rain legislation. Ground water runoff and contamination is another area where future legislation would seem likely. Already, one bill has been introduced in Congress. A second is being drafted. The impact of such legislation may be significant according to Bruce Leavitt, a hydrogeologist with Consolidation Coal Co. He said if current proposals are adopted, there will be more federal, state, and local government involvement in ground water regulation. In any event, the coal industry can expect to see more emphasis on preventing acid mine drainage and on water replacement, according to Leavitt. He urged those in the coal industry to present information about mining and ground water. That is needed to prevent misdirected state and federal programs, he said. Another coal industry concern is excess capacity. The industry has the mines, equipment, and employees to produce 15% more coal than at present. Problem is, the markets are not there. Slower-than-predicted growth in electric utility coal use has kept sales sluggish. There are also tax uncertainties. Congress is considering repeal of the investment tax credit and elimination of black lung payments and excise taxes as deductible expenses. One analyst estimates the coal industry would lose $1.1 billion in five years, if the changes are approved. In addition, there are the usual concerns about excessive governmental regulations involving safety and environmental matters. Bill Kegel, for example, said these regulations mean extra costs and delays in developing mines. Kegel is president and chief executive officer of the Rochester & Pittsburgh Coal Co. More than half the electrical power in the US is generated by coal-fired plants. That percentage could slip by a couple of points as nuclear generators come on-line the next few years. About 1990, though, we will see the end of US nuclear plant construction. No new nuclear plants have been scheduled since 1978. So any growth in electric power use should benefit the coal industry. BethEnergy - High Power Mountain During 1985, BethEnergy - a Bethlehem Steel Corp. - subsidiary developed High Power Mountain, a 1.8-Mt/a (2-million-stpy) surface mine in West Virginia. Construction saw movement of more than 3 hm3 (4 million cu yds) of earth. A computerized 544 t/h (600 stph) heavy media cyclone prep plant and a 3.6-kt/h (4000-stph) railroad loadout facility were built in six months. And a 5.6-km (3.5-mile) railroad spur and loop bridging a major highway were constructed. Larry Willison of BethEnergy noted the project's ambitious construction schedule. It was forced by the need for the project to be market driven and - lacking available capital - externally financed. BethEnergy did several things before obtaining with Detroit Edison a market for 0.9 Mt/a (1 million stpy) of coal. Willison said his company prospected and proved the eastern half of its 8-km2

Jan 7, 1986

By Hans Ehdwall

INTRODUCTION Investigations of radon and radon daughter concentrations in Swedish [non-uranium] mines started in the late 1960's. The first screening measurements showed that the average annual exposure to radon and radon daughter products was 4.7 WLM. The main reason for high radon and radon daughter concentrations was inefficient ventilation and radonrich water entering the mine. In the radon regulations worked out later it was stated that no miner should be exposed to more than 60 000 pCi h/1 equilibrium equivalent concentration of radon annual exposure, corresponding to 3.6 WLM. Now, 1981 the situation has changed considerably. From the average annual exposure of 4.7 WLM in 1970 it is now only 0.7 WLM. Sweden has up to now had only one [uranium] mine and the work there has only been investigative. However, there are plans for a commercial uranium mine in another part of Sweden. The radon problems in these mines are widely different depending on the mineralogy. NON-URANIUM MINES The radiation problems in Swedish mines were not recognised until the late 60's. The first radon and radon daughter measurements were made in some sulphide ore mines in 1967 (1). The radon and radon daughter concentrations were surprisingly high for non-uranium mines. In order to have a complete picture of the radon situation in Swedish mines the National Institute of Radiation Protection (NIRP) decided to make measurements in all, at that time about 60 mines (2). To get results as fast as possible measurements on radon gas seemed most appropriate to start with. Sampling was made by mailing a number of evacuated 4.8 litre conventional propane containers from NIRP to each mine. The containers were then opened at the place of interest. After sampling the containers were sealed and then mailed back to the institute for measurement. The measurements were made in ionization chambers. This method only gave the radon concentration and the radon daughter concentration was estimated by multiplying the radon concentration by an assumed equilibrium factor. The equilibrium factor is defined as the ratio of the total potential alpha energy for the given daughter concentration to the total potential alpha energy of the daughters if they are in equilibrium with the given radon concentration. The results of this first preliminary survey indicated that a great many of the Swedish miners probably had an annual radon daughter exposure of more than 3.6 WLM. As the radiation exposure in non-uranium mines was not regulated in either the Swedish Radiation Protection Act or the Swedish Labour Protection Act work was started on special radon regulations. A lung cancer mortality study was also started. To check the results of the first survey and to get experience and knowledge of radon problems in mines, it was decided that personnel from the NIRP should visit each mine for a detailed investigation of radon and radon daughter concentrations starting with the ones with the highest radon concentrations. The main reasons for these so-called "basic measurements" were: 1. To estimate the doses received by Swedish miners 2. To find the sources of the high radon and radon daughter concentrations 3. To find appropriate counter-measures 4. To determine the most typical equilibrium factor for each mine. Unlike most uranium mines the reason for high radon concentrations in non-uranium mines is seldom the occurrence of highly radioactive minerals. The main sources were found to be waste-rock and radon-rich water. In order to filter and warm up the inlet air, especially in winter time, it was very common at that time to suck the air through broken wasterock. By doing so the air was contaminated with radon from the waste-rock and radon-rich water in it. It is noteworthy that the radium and uranium concentration in the waste-rock is relatively low. The uranium concentration is only of the order of 15 - 20 ppm. The action to prevent this contamination of the inlet air was to change the direction of the ventilation and in the case of radon-rich water entering the mine the action was to prevent the air coming into contact with the water. The first calculation of the radon daughter exposure of Swedish miners was based on radon gas measurements. The radon daughter concentration was estimated by using an assumed equilibrium factor of 0.5. Later when the mines were visited by institute staff it was possible to compare the assumed equilibrium factor with the measured ones. It was found that the factor varied from 0.15 at the air inlet to 1.0 at the air outlet and the average equilibrium factor on workplaces for almost all mines was between 0.4 and 0.6. The result of the exposure calculation in 1970 showed that more than 40 % of the miners had an annual radon daughter exposure of more than 3.6 WLM. The overall average was 4.7 WLM and the maximum annual expo-

Jan 1, 1981

By Ireneusz Tomza, Jolanta Lebecka

INTRODUCTION Radioactive deposits were observed in 1972 in some of the Upper Silesian coal mines. They were located mainly in the drains in galeries and on the inside surfaces of water pipes. They also caused some problems by accumulating in water pumps. It has been postulated that the deposits are produced by natural radioactive waters seeping from the rocks. Investigations were initiated to answer the following questions: - What is the composition and the amount of radioactivity in the deposits? - What radioisotopes are present in the water? - How are the radioactive deposits formed? - Do the radioactive waters also occur in other mines? - How does the radioactivity of the water depend on chemical composition? - What is the origin of the radioactive water? - Does the water and the deposits cause radiation hazards for miners? -How can the radiation hazard be reduced? METHODS OF MEASUREMENT Determination of Radium Isotopes in Water The commonly used methods of radium determination in water are either based on measurements of the radioactivity of 222Rn which is in equilibrium with 226Ra, or on the detection of alpha particles of the radium radioisotopes after chemical separation of radium from the water sample. The method based on radon activity measurements is very sensitive and does not require any chemical Separation, but it can be used for determination of 226Ra from the uranium series only, because the thorium daughter 220Rn has too short a half-life (55s to yield the required accuracy. The method developed by Goldin, 1961 [2] involved alpha-particle measurements in thin layers of RaS04 and BaS04 separated from the water. This method is not convenient for saline water and water with high barium concentration because the amount of barium carrier in this case is too large to obtain a thin layer of precipitate with sufficient activity. The Upper Silesian carboniferous waters are often saline with high barium content, so the method described by Goldin was not convenient for this case and it was necessary to change the detection system and modify the chemical preparation. The procedure developed by the authors for the determination of radium isotopes in water was as follows: - Depending on the Ba2+ content and the required sensitivity of measurement, a water sample of 200 cm3 to 3 dm3 was taken. - 10 cm3 of 0.25 M citric acid and 5 cm3 15M ammonia was added to form complex Ba2+ ions and avoid the immediate precipitation of BaSO4. (This was repeated as long as the addition of BaC12 did not form a precipitate.) - 1 cm3 of 1N solution of Pb(N03)2 as a carrier for radioactive isotopes of lead and 10 cm3 of 0.1 N BaC12 as a carrier for radium were added. - The sample was heated to the boiling point and the precipitation of RaS04, BaSO4 and PbS04 with 50% H2SO4 was carried out. - After several hours the sample was centrifuged and the precipitate was purified by washing with nitric acid and distilled water. - The precipitate was then redisolved in 20 cm3 0.125 M Na2EDTA and 3 cm3 6M ammonia and reprecipitated from the solution by dropwise addition of acetic acid to d pH of 4.5. At this value of pH, precipitation occurs only for the barium and radium sulfates, while lead and all other radioactive elements remain in the solution. The date and time of deposit precipitation was recorded. - The final barium-radium sulfate mixture was washed with distilled water and transferred to standard measurement vials. - Each vial containing a deposit had 6 cm3 of distilled water added and was then shaken vigorously. 12 cm3 of liquid gelling scintillator (INSTA-GEL UNISOLV-1 type) was then added and the vi 1 was shaken again. After a while the scintillator turns into a milky gel in which the deposit is uniformly distributed. - The standard sample of 226Ra was prepared in the same way. - The activity of the samples was measured using a liquid scintillation spectrometer. (In this case the TRICARB 3320 produced by Packard Instruments, was used). Tests run on standard radium solutions provided by Amersham Radiochemical Centre indicated that this method of measurement enables one to achieve an efficiency of almost 100% (within measurement error). For alpha particles no quenching effect was observed for the BaS04 concentration in the range up to 80 mg of BaS04 per 1 cm3 of liquid scintillator coctail (Fig. 1). This provides a sensitive determination of radium in water with high barium content and also in saline water. In saline water the solubility of barium sulfate is much higher than in

Jan 1, 1981

By A. J. Kroha, N. Wesis

Most copper mines produce both ore and low-grade "leach" rock or acid waters that contain recoverable copper. The sulfide ores pre¬dominate, and a portion that is too low grade for milling to produce concentrates for smelting, but has to be mined and trucked away anyhow, may be leached successfully with acid in dumps. Most of this leach material consists of sulfides and silicates or carbonates, and if the gangue is such that it consumes a high quantity of acid, this factor may rule out a leach operation. There are also valuable deposits that contain mostly acid-soluble copper, or occasional sulfide ores from which a sulfide concentrate can be roasted and acid-leached to produce a copper-bearing solution. Finally, there are milling ores in which the lesser part of the copper is acid-soluble and can be precipitated with iron or synthetic inorganic precipitants that produce metallic copper or copper sulfides that will float with the sulfides. Ordinarily, ores that contain copper associated with the sulfur ion, such as in the minerals chalcopyrite, chalcocite, bornite (and others), are milled to produce a 25-30% Cu concentrate for smelting, while a lesser amount of acid-soluble copper may be converted from solution to cement copper on iron scrap. A fast-growing percentage of such copper, however, is removed from solution with exchange resins or organic compounds in organic carriers such as kerosene (solvent extraction), then eluted with strong acid and subjected to electrolytic precipitation either in marketable form or as anodes that can be refined further. From the point of view of conventional copper smelting, copper flotation concentrates and cement copper are of chief interest in this chapter. Table I is a condensed open schedule for concentrates that generally run between 25 and 35% copper, and much less frequently as low as 12-15% or as high as 65-75% copper, the former being due to intimate relationship with pyrite (like the former United Verde Extension), and the latter representing such ores as the Bolivian Coro¬coro ore in which the copper is in the form of chalcocite in sandstone. These extremes are no longer common. When they occur, a special purchase schedule has to be negotiated. Included in Table 1, copper precipitates (cement copper) generally run from 70-85%a copper, and the same basic purchase schedule is used as with flotation concentrates. Sulfide Flotation Concentrates The sulfide copper concentrate produced in the mill as a flotation froth, with water then added for transportation of the heavy mineral particles from the flotation cells to thickeners, may run 60-80% water by weight, and the removal of water down to 25-50% by weight by means of thickeners, followed by further dewatering by continuous vacuum filters to 7-18% moisture by weight (depending on size of solids by screen analysis and also by content of clay) is a critical operation. Mill operators would like to produce a filter cake with 7-9% moisture content, but even with the help of steam on the filter this desirable condition is seldom realized when the concentrate is as fine as 80% -325 mesh. More commonly, the final concentrate is reground in pro- to produce best copper recovery and grade of concentrate (or molybdenite separation). In those cases, increasingly frequent, the filter product may not be a cake at all, but a mud that is hard to handle-even requiring a thermal dryer. Greater difficulty of form¬ing a manageable cake often comes from the copper-molybdenum separation by flotation, because the alkaline sulfides and hydrosulfides, or cyanide, or other similar reagents used for the separation, may leave the now relatively molybdenite-free copper concentrate even more difficult to filter. Handling a wet filter cake is difficult enough when its destination is only a short distance away-a matter of yards rather than miles. In those cases the filter cake may be thermally dried near the point of production, using rotary or multiple hearth, or fluidized-bed dryers. Alternatively, the concentrate may be pumped or carried in slurry form to the smelter and filtered there, or it may be spray-dried and compacted. For transportation to a smelter just a few miles to a few thousand miles away by ship or railroad other factors may be important, such as: in shipping by sea, avoidance of spontaneous combustion; in shipping by rail, losses by leakage if too wet or by wind and sun if too dry. It is the responsibility of the millman-usually the mill superinten¬dent-to make sure that his concentrates are in satisfactory condition when they leave the mill so that they meet these requirements: 1) They must have been accurately sampled and dry-weighed, the latter meaning that a moisture determination and gross weight must have been taken. 2) They must be dried sufficiently when necessary to prepare them for safe transportation. 3) They must arrive at the smelter with reasonable likelihood that they can be check-weighed and sampled fairly and equitably,

Jan 1, 1985

By Martha E. Smith

INTRODUCTION Radiation is one of the many agents occurring in our environment that is capable of causing cancers. We are all unavoidably exposed to natural background radiation. If we wish to exploit the beneficial uses of radiation in medicine, industry, and for the production of electricity, some further exposures to radiation are almost inevitable. One needs to ensure therefore, that the standards of radiation protection are safely derived and one would want to understand the mechanism by which radiation might cause cancer and genetic defects. In order to study the effects of radiation one can: (1) analyze human data, (2) perform animal experiments, and (3) do molecular level experiments. Indeed all three kinds of information are needed. This paper will concentrate on the steps required to obtain some of the necessary human data regarding health effects, where the concern is about the impact and consequences of long-term low-level exposures to an agent. Such investigations require some knowledge of work histories, dose histories, health "outcomes", and the personal identification of the individual involved. Three inter-related computer systems have been developed at Statistics Canada which have been designed to permit optimal use of a number of different records for our entire country for such health related research. The development of the Canadian Mortality Data Base, the initiation of the National Cancer Incidence Reporting System, and the development of new computer linkage techniques have helped reduce the cost and increase the scale and efficiency of automated follow-up to produce statistics of sickness or death associated with radiation and other carcinogenic agents in mines. These computer systems have already been implemented, and references will be made to studies currently being conducted using these files and facilities (e.g. a study of all Ontario miners, plus various Canadian uranium, fluorospar, salt and nickel miners). We will also look at the kinds of data that need to be collected now, to improve such studies in the future. DELAYED RISKS - THE STUDY SIZE AND COST Delayed effects on human health, as for example industrially caused cancer, can in general only be detected and measured by following-up the individuals to see what eventually becomes of them. What is not generally recognized is that the relatively low levels of individual risk, about which the public is often concerned, usually requires for their detection that very large numbers of "exposed" and "control" individuals (e.g. 10,000 to 100,000 or more) be followed over a period of two or three decades to determine when they die, what they die of, and whether they contracted cancer or some other disease of special concern. Thus, it is frequently exceedingly difficult to make such investigations cost-effective so that they will be undertaken at all, and as a result very real risks to health can remain undetected or unquantified. THE MANUAL PROCEDURES Follow-up of individuals by epidemiologists has until recently been a largely manual and clerical operation. It has used a diversity of source record files; local, regional, and national. Often the tracing of people has involved letters sent through the mails and visits to institutions, physicians, municipal offices, and former neighbours. Only thus could one find out whether the individuals were dead or alive. Such studies were necessarily small, or else very expensive. Death registrations have, by tradition, provided a valuable tool for the identification of harmful influences in the environment. However, with the increasing mobility in the population, death may often occur far away from the place of exposure to such an influence. Thus, no longer will simple manual searches in a single registry office suffice to inform the investigator concerning the deaths that may have occurred in a study population, especially when that population is large. In the past, manual follow-up to locate the relevant death registration normally required some prior knowledge of the province and year of death in question, so that the alphabetic indexes could he used to direct researchers to the appropriate bound volumes of registration forms. To search manually in this fashion for any large number of death registrations, without knowing in which year or province the deaths had taken place, or even whether they had yet occurred, would be impractical, since each year in Canada there are about 170,000 deaths. DEATH AND CANCER AS SPECIAL ENDPOINTS Much of recent effort has been focused on the organization of the "endpoint" files required to do long-term follow-up studies on a national scale, because that is a function which other institutions are unable to perform due to the confidentiality laws governing the use of such information. Outside organizations generally come to us with detailed "starting" point records which relate to some specific group requiring study. We carry out these epidemiological searches on a cost-recovery basis. The analytical interpretation of results is normally

Jan 1, 1981

By R. Masse, J. Chameaud, J. Lafuma, R. Perraud

Estimating the risk of lung cancers for workers in uranium mines and defining the resulting dose equivalent limits have been made possible thanks to work carried out in two scientific fields : physics and epidemiology. Theoretical calculations on the basis of physical models for the former and epidemiological surveys on the mortality of uranium miners from lung cancer for the latter. However, even though considerable work has been done in these two areas, the results obtained still remain controversial on several points. The radioactive and physical instability of the aerosols present in the atmosphere of the mines and the biological complexity of the human lung which even the most sophisticated physical models can reproduce only very schematically have often proved to be insurmountable difficulties for physicists, explaining the uncertainties which subsist concerning the dose delivered to the various parts of the respiratory tract from the air breathed by miners during their work. Epidemiological investigations on the other hand, in spite of the high quality of the surveys carried out, remain open to criticism, essentially because of the very approximative estimation of the individual occupational exposure to radon daughters. This is due to the fact that uncertainties arise from the measurement of radon gas if the state of equilibrium with the daughters is not accurately known or, if the active deposit is measured, to the fact that these measurements are insufficient in number. The controversies and discrepancies which subsist with regard to the evaluation of the level of risk, and in particular for low doses, can thus be understood. In addition, epidemiological surveys cannot dissociate the carcinogenic action of radon from the synergistic or potentiating actions of tabacco and of other pollutants present in uranium mine air. Animal experiments have been largely taken into account for evaluating the toxicity of various radionuclides. This type of experiment is necessary when human data do not exist and has provided us with much information. For instance, the relative biological effectiveness of the various types of radiation, the metabolism of radionuclides and the mechanisms of cancer induction have been approached and satisfactorily resolved in this way. Concerning radon and its daughters, however, animal experiments have been used very little even though it seems apparent that they should complement epidemiological studies. For instance, whereas doubt can be cast on the data obtained from human epidemiology because of the uncertainty concerning the individual exposure of miners, those drawn from experiments are indisputable because in this case the dose is as perfectly known as the effect. In addition, the effects of radon can experimentally be appreciated separately whereas in the surveys, they cannot be dissociated from the effects of the other pollutants in the mine. Finally, there are no other means of dealing with the mechanisms of cancer induction. In order to gain any useful knowledge from this method however, the experimental model must necessarily present certain methodological guarantees and the effects seen in the animals must enable a comparison with those which appear in man. For this reason we will present here the animal model we have been using for 15 years, and will give the results obtained and compare them with human data and made a synthesis. Finally the conclusions which can be drawn will be discussed as well as their limitations with respect to the protection of uranium miners. I - MATERIAL AND METHODS Male SPF Sprague-Dawley rats were used. At the onset of the inhalations they were around 3 months old. Their small size makes it possible to expose a large number of animals at the same time. Their life-span is long enough to be able to follow the evolution of the cancers and to estimate the latency time. Finally, they present the advantage of having a very low rate of spontaneous lung cancers (SANDERS, 1979). I.1 - Three inhalation techniques were used. 1.1.1. – [Inhalation of radon decay products.] The inhalation apparatus has been described previously (CHAMEAUD et al. 1971). The first experiments utilized a room of a half cubic meter linked to a source made up of high grade uranium ore. Later on, a large installation was built with a 10 m3 inhalation chamber making it possible to expose up to 500 rats at one time at radon concentrations ranging from 100 to 10 000 WL for variable lengths of time (1 to 10 hours per day). These concentrations are higher than those to which the miners are generally exposed, but in order for the cumulated doses in man and in animal to be similar and delivered for the same fraction of their respective life-spans, the ratio of the concentrations should be approximately that of the life-spans. The concentrations of radon and its daughters during the experiments were carefully controlled thanks to multiple samplings of radon gas associated with measurements of radon decay products. I.1.2 – [ The dust inhalation chamber] has already been described : it is a dust-loading chamber where the dust content remains constant during the experiment and can hold over 20 - 30 animals (PERRAUD et al, 1970). I.1.3 – [Tobacco inhalations] take place in a smoke box

Jan 1, 1981

By J. Duncan Wilkins

Introduction In difficult times such as these, there is a strong reaction to the current way of doing things. Typical reactions that we have all heard are "There has to be a better way," "We're pricing ourselves out of business," "We have to improve our productivity," and "We have to have more cooperation," between union and management and employees and management. All of these comments have at their roots one common factor - getting the maximum amount for your dollar. The other factor inherent in these statements is productivity. I have not yet met a person in our industry who has not expressed the opinion that we should, and that we can, improve our productivity. Reducing Labor Costs In light of these factors I suspect we have all spent a fair amount of time examining our labor costs. For some of us, labor costs are a high proportion of our total cost of doing business. To alleviate the impact of these costs, we have generally done four things. We have reduced our number of employees, shut down operations for appropriate periods, sought concessions from union employees, and placed freezes on wages, salaries, and other benefits on nonunion employees. These approaches have been made to improve current shortfalls in our cash positions - to tide us over, as it were - or to provide us with a chance for survival. These are short-term measures that help to bring immediate relief, but can pose significant problems (or challenges), for the longer term. For instance, what do we do when times improve? How much do things have to improve before we do anything? By seeking concessions from unions in bad times, what do we do when unions come to us in good times? It is a rather sad and critical fact that we have grown too fat during the good times and too thin during the bad. In the first case, we have failed to optimize our earnings. In the second, we have cut ourselves too far. Consequently, when good or better times have arrived, we have had to bulk up our requirements to meet production commitments. In mining, for example, when times become tough, we tend to reduce our development plans so that when times improve we have to really "sock it to 'em," so that we can maintain productive capacities. We should plan ahead a little more to reduce the amplitude of our cyclical wavelength, so that in good times profits are optimized and in bad time we are better able to take the strain. Of course, forecasting cycles is not a refined art, but if we properly control our work force levels and costs at all times, and therefore optimize our productivity, we would be more able to withstand the problems we face today. Unfortunately, it appears that it takes bad times to bring us to a realistic appraisal of our way of managing our businesses. Labor Practices Not all of the problems now faced are due to low metal prices. Inflation has played a major role in bringing costs to a frightening level. Trade unions alone can not be blamed for high inflation levels over the past several years, popular though that notion is. Contract negotiations, after all, require two parties. Indeed, our current levels of labor cost are due to two factors: • We have felt obliged to keep our employees whole, relative to the cost of living. • We have felt obliged to maintain our competitive position relative to our peers, in order to maintain our ability to attract and retain a skilled, efficient work force. In the first case, our felt obligation has been applied without due recognition of the factor of performance, either in individuals or groups. Average, even mediocre performers, have been amply rewarded for their average work and mediocrity, while the good performers, no doubt receiving more for their good work and effort, have not perhaps appreciated the slight premium for their effectiveness. The rather predictable result of this practice has been that the average and mediocre stay that way (why change?), while some (certainly not all) of the good performers have said "It's not worth it," and slipped into the warm, cosy pool of the average and mediocre. In the second case, that of maintaining competitiveness with other companies' employees, we have compounded an already serious problem of lack of skilled tradesmen by paying higher and higher prices for the existing pool to maintain comparability without improving the flow of more skilled people through sound and sufficient training programs. We have the rather dubious pleasure of paying more and more for the same problems. Compensation is either mutually agreed, as with unions, or it is unilaterally applied, as with nonunion employees. Generally, there is more flexibility available to the employer when dealing with nonunionized groups than there is with unionized groups. Sometimes that flexibility has not been used well, most often be-

Jan 11, 1983

By Jan Tegtmeier, Horst Gerhardt

INTRODUCTION Under certain circumstances the closure of former mines which are located above a certain flood level can result in problems such as the emanation of detrimental substances after having completed filling and reclamation operations. This especially applies to uranium mines in which the radiation dose could far exceed the dose of natural background radiation. By means of an example of the uranium mining in Germany in the following it will be demonstrated how to cope with this problem. On the basis of comparative investigations in various vein deposits and using ventilation scheme calculations proposals for the optimization of the necessary forced ventilation can be submitted. REPORT ON SITUATION In the period 1946 - 1989 the former Soviet-German joint- stock company "Wismut" developed into the biggest European uranium producer with a total output of about 220.000 t of uranium. A major mineraldeposit district was the deposit of Schlemaf Alberoda in the Saxon Ore Mountains, in which 80.000 t of uranium were produced. Thus it is among the biggest uranium de- posits of the world, from which various other metals were at- tracted for many centuries. The exploitation of the Schlemal Alberoda deposit involved steep veins in regions near the surface as well as depths of 1.800 m. Until 1991 a total excavation space of 40 million m3, which is flooded at present, was produced. With the average increase in the water level of 80 cm per week the final flood level is expected to be reached in the year 2003. The shaft 373 at present still being used for ventilation will be no longer available since the second quarter of 1998 after flooding the -540 m level because it is not connected with the excavation system near the surface. As a study shows, a radiation dose far above the natural back- ground radiation has to be expected for the town of Schlema due to the extensive mining activities near the surface and due to the subsequent displacement with missing depression fo the main mine ventilating fan. An uncontrolled air flow containing radon leaves the open mine excavation due to the effect of the natural ventilating pressure and emanation caused by the barometric pressure drop with atmospheric pressure fluctuations. This mine air with its high-level radioactive equilibrium results in a high radiation dose in buildings (see Figure l). After having switched off the main ventilating fan in order to investigate the effect of the missing depression the increase in radon concentrations amounted up to 700% in various buildings of Schlema. This was partially due to the inversion state of the weather at that time. The high radon concentration has detrimental effects on the health of the population and of the miners working on the further reclamation in regions above the flood level. ANALYSIS OF THE RADON EMANATION RATE EXPECTED Considering the composition of the radon inflow from the mine workings it becomes evident that 80 % of the radon inflow originates from abandoned excavations and only 20 %from open ventilated mine excavations. This fact has to be taken into account for the ventilation after having reached the final state of flooding. After completing ventilation the radiation dose on the surface is mainly due to the radon emanation from excavations close to the surface. Investigations of the Wismut GmbH showed the in- crease in the specific radon emanation rate by a factor of 100 for abandoned excavations as compared to new drivings. One reason is the larger specific surface of abandoned galleries caused by displacements due to mining activities as well as by fall of hanging. Furthermore the radon can enter the gallery through joints, which have subsequently opened by convergences. All these effects result in a larger free surface available for radon diffusion. The large number of drivings in the deposit sections near the surface and the fact that the highest uranium contents are found near the surface as well as the high fracturing are further reasons for higher emanation rates. Considering these facts it can be expected that the radon inflow of 10.000 kBq/s, which refers to an open mine excavation of about 1.4 million m3, represents a minimum. Only by increasing the specific surface, for which a numerical value has still to be determined, this value will increase with certainty. An extensive radon emanation from the residual excavation, which cannot be flooded, can only be prevented by maintaining the ventilation system. The low pressure produced by the fan in the mine openings prevents the emanation of air containing radon due to the effect of the natural ventilating pressure. Without the controlled withdrawal of the radon the population as well as the miners working on the further reclamation in areas above the flood level would be endangered. Therefore the follow-

Jan 1, 1996

By B. P. Boisen

INTRODUCTION The rapidity of onset, rate of increase, and magni¬tude of loads in an underground support system can be measured using load cells or pressure cells, whichever is appropriate to the type of support. These instruments should be installed at the various instrumented station locations immediately after excavation. If the loading of a member is of interest, the load should be measured, whereas if the deflection or strain of a member is of interest, the strain should be mea¬sured. It makes as little sense to use a load cell to allow computation of strain as it does to measure the strain and then back calculate the load. Tunnel steel set de¬signers today invariably rely on the early work of Karl Terzaghi, the basis for which is load, so load measure¬ment should be the main concern. The current trend of using strain gages on steel sup¬port systems stems from the inability by some to eval¬uate unusual loading conditions caused by uneven block¬ing. In fact, uneven blocking will obscure almost any attempt (whether with load cells or strain gages) to properly evaluate support performance. The load curve in Fig. I shows the development of a characteristic early peak load sometimes called the "rear abutment load" seen in many underground open¬ings. It is thought that the peak reflects the coupling of the rock and the support system, and that the magnitude will increase until some yielding, or minor failure, of the support system occurs. At that time, the slight deforma¬tion of the support system promotes the formation of minute shears in the opening walls, and these shears tend to distribute stress between the support system and the adjacent rock in proportion to the relative rigidity of the support elements and the rock. In normal rock, as a result of this stress redistribu¬tion, subsequent load magnitudes generally do not reach the magnitude of the early peak load. In physically unstable (squeezing) or chemically un¬stable (swelling) rock, however, the loads experienced after passage of the early peak load may in fact show a slow, continuous increase. One of the very important purposes of load instrumentation is to provide the means for recognizing such long-term adverse trends, thus enabling the proper remedial steps to be taken. Another purpose of load instrumentation is to pro¬vide a comparison between the magnitudes of the early peak load and the subsequent stable load. The ratio of these two loads is analogous to a safety factor, and may be used to evaluate the efficiency and economy of the support system design. Aside from considerations of economy, it may be well to design support systems which do not have excessively high ratios of early peak load to subsequent stable load. Should these ratios be exces¬sively high, the support system may be so rigid that the yield or failure associated with stress redistribution may occur with explosive violence. Load cells for use in mines, tunnels, and on con¬struction projects come in many forms. Almost all, however, employ the same procedures for installation, readout, etc. Therefore the following comments are almost universal in application. LOAD CELL INSTALLATION Load cells should be installed with bases parallel to the surfaces against which they bear. Care must be taken to orient the cells so that their signal cables are protected from accidental damage as a result of con¬struction, maintenance, or cleanup activities. Most electronic load cells are compensated for tem¬perature variations likely to be encountered during nor¬mal operations. However, if a large difference is an¬ticipated between the calibration temperature [21°C (70°F)] and the average operating temperature, the cells should be conditioned to the operating temperature for at least 8 hr prior to installation. This is to insure that the initial reading, made under no-load conditions prior to installation, provides a stable value to which subse¬quent measurements can be referred. Hydraulic load cells tend to be temperature sensitive and should be used with that in mind. Also, hydraulic load cells tend to be soft compared to electronic types and will sometimes allow movement to take place in a system intended to be semirigid. Furthermore, hydrau¬lic load cells tend to be difficult to read remotely. Care must be taken to use bearing plates on both sides of load cells which are sufficiently rigid and of high enough bearing capacity to prevent bending and crush¬ing under load. This is very important with tieback load cells (basically a ring of steel) which can easily dig into bearing plates. MAINTENANCE AND TROUBLESHOOTING Except for a direct hit by a miner's axe or flyrock from a blast, most hydraulic load cells are nearly in¬destructible and require little, if any, maintenance. Hy¬draulic oil on the bearing plates is a good indication of a leak and the need for corrective action. Field maintenance of electronic load cells involves protecting the instrument and signal cable from mechani¬cal damage and from unnecessary exposure to dirt and moisture, and recognizing and correcting damage and the effects of normal wear and tear.

Jan 1, 1982

By Erhard M. Winkler

The rapid decay and disfiguring of stone monuments in urban and desert rural areas has challenged conservators to protect stone surfaces from premature decay. They attempt to halt the natural process of stone decay and possibly to restore the original strength lost mostly by chemical weathering and the loss of binding cement. Ageneral solution is not possible because the physical and chemical characteristics must be considered for different stone types. The failures of stone preservation and restoration are greater in number than the cures. The need for repair of stone decay goes back to evidence of Roman replacement of decaying stone. The presence of excess water in buildings has long been recognized. Moisture tends to enter masonry from air in humid climates, a most important but often underrated factor (Fig. 1) suggesting that sealing should be the answer. Undesirable staining and efflorescence result in accelerated scaling. Today, the great variety of chemicals available to the modem conservator for sealing. consolidating, or hardening stone fall into two very different categories: surface sealers and penetrating stone consolidants, or a combination of both. SEALERS Sealers develop a tight, impervious skin which prevents access of moisture. Surface sealing has saved monuments from decay by eliminating the access of atmospheric humidity. Pressure tends to develop behind the stone surface by moisture escape. Efflorescence, crystal growth action, and freezing can cause considerable spalling (Anderegg, 1949). Flaking results when moisture is trapped behind the sealed surface. Yellowing and blotchiness are also frequently observed. The following sealants are in common use today: linseed oil, paraffin, silicone, urethane, acrylate, and animal blood on stone and adobe. Extensive cracking and yellowing has resulted soon after application. In the past many such treatments have created more problems than cures: 1. Linseed oil and paraffin have been in use for centuries. Embrittlement and yellowing occur rapidly because these are readily attacked by solar ultraviolet radiation. 2. Animal blood as paint has temporarily waterproofed adobe mud and stone masonry. The origin of blood paint has a religious background rooted in the Phoenician and Hebrew cultures. Instant water soluble dried blood can substitute for fresh blood. Winkler (1956) described the history and technique of the use of blood. 3. Silicones have proven very effective and are long lasting. In contrast, acrylates, urethane, and styrene are generally rapidly attacked by UV radiation (Clark et al., 1975). Sealing of Different Rock Types Granitic rocks have a natural porosity traced to 4.5% contraction of quartz, during cooling of the parent magma, compared with only 2% contraction of all other minerals; protection against the hygric forces may require waterproofing of granite in some in- stances. The Egyptian granite obelisk in London is an example. Soon after its relocation from Egypt to London, Cleopatra's Needle was treated, in 1879, with a mixture of Damar resin and wax dissolved in clear petroleum spirit; surface scaling became evident after half a year of exposure to the humid London atmosphere. The treatment of the ancient granite monument from Egypt has denied access of high relative humidity (RH) in London to the trapped salts inherited from the Egyptian desert and has protected the monument from decay (Burgess and Schaffer, 1952). The sister obelisk set up in Central Park, New York City, has fared less favorably because similar treatment was done too late, only after the salts hydrated and hundreds of kilograms of scalings disfigured the obelisk surface (Winkler, 1980). Surface coating of other common stones may be needed. Crystalline marble absorbs moisture from high RH atmospheres: dilation may ensue when curtain panels bow as the moisture starts to expand during daily heating-cooling cycles. A good sealer may prevent the moisture influx provided that no moisture can enter from the inside of the building. Limestones, dolomites and all carbonate rocks are subject to dissolution attack by rainwater, especially in areas where acid rain prevails (Fig. 2). The interaction of sulfates in the atmosphere with the stone can be halted by waterproofing to avoid the formation of soft and more soluble gypsum. The stone surface attack can be diminished if nearly insoluble Ca-sulfite crusts can form, instead of Ca-sulfate. Replacement of fluorite or barium compounds at the stone surface acts as a hardener, rather than a sealant. Sandstones have generally high porosity and rapid water travel can occur along unexpected routes and from any direction. Any surface sealing may do more damage by scaling and bursting than if the stone is left without treatment. Sealing of sandstones is therefore not advised at any time. Testing the efficiency of sealants: Several authors discuss waterproofing materials, silicones, urethanes, acrylates and stearates, as to their water absorption, spreading rates of water on the treated surface, water vapor transmission, resistance to efflorescence, and general appearance (Clark et al., 1975). De Castro (1983) measured the angle of contact of a microdrop (0.004 cm3) on a stone surface as characteristic of the wettability. Laboratory tests and limited field performance are described by Heiman (1981). The crest of a Gothic sandstone arch, which was sealed with silicone,

Jan 1, 1994

By Christopher M. St. John, Michael P. Hardy

INTRODUCTION Geotechnical models, particularly those based on the finite element method, have been available to aid in en¬gineering design of underground mining excavations for over ten years. Despite this fact there are remarkably few cases of their use in mine design documented in the literature. It is therefore to be anticipated that many potential users of these models are relatively unaware of their capabilities and limitations and also of the form and detail of geotechnical data needed for their success¬ful application. This subsection attempts to address this problem by discussing the different types of numerical models now available and by noting how some of them have been used to study a variety of problems associated with underground mining. The subsection concludes by discussing the applica¬tion of the displacement discontinuity method to the de¬sign of possible mining systems for the copper-nickel deposits of northeastern Minnesota. The object of the analyses, which were nonsite specific, was to determine the significance of geotechnical parameters, such as ini¬tial stress and rock structure, on the stability of under¬ground excavations and hence to provide guidance for future geotechnical investigation. NUMERICAL MODELS In the geomechanical design of underground mining openings, use is made of numerical models that repre¬sent or simulate the large-scale mechanical behavior of rock. There has been less interest in analysis involving fluid flow and heat transfer, but with increasing interest in such areas as in-situ retorting and solution mining it is likely that there will be a growing need for numerical models embracing and coupling all three physical proc¬esses. However, the emphasis in this subsection will be on mechanical behavior. Models which simulate such behavior will be divided into two groups: continuum models and discontinuum models. These will be dis¬cussed in turn in order that some insight into alternative solution strategies and their merits may be gained. Continuum Models Almost all geomechanical numerical models must be classed as continuum models even though particular computer codes incorporate special provisions for rep¬resenting discontinuities such as faults, bedding planes, or joints. They are continuum models because they pro¬vide solutions for cases where material behavior is governed by the differential equations of continuum mechanics. Two basic solution strategies for such equa¬tions may be identified immediately: the differential ap¬proach and the integral approach. In the differential approach a means of approximating the differential equations over the entire region of interest is sought. In the integral approach use is made of fundamental solutions from continuum mechanics, and these are used to construct a solution to the whole problem, making approximations only on the boundaries of the region of interest. The several differential and integral methods are identified in Table 1. Differential Methods: Problems in continuum me¬chanics involve the solution of three types of partial differential equations. Two of these govern the behavior in so-called initial value problems, in which variables change both in time and in space. Examples of such problems include nonsteady heat transfer and fluid flow, and stress wave propagation. The last type of partial differential equation governs the behavior in boundary value problems. In these, variation is in space but not in time. Solution of initial value problems may be achieved in two significantly different methods: implicit and ex¬plicit. The differences between these two methods will be illustrated by considering a very simple initial value problem, that of one-dimensional heat diffusion. The equation governing this process may be written as: [ ] where T is temperature, K is thermal conductivity, p is density, c is specific heat, t is time, and x is the spacial coordinate. In finite difference form this equation might be written as: [ ] where the superscripts refer to the time and the sub¬scripts to the spacial location. Several solution strategies for this equation have been used. Two important ones may be illustrated very simply by discussing the signifi¬cance of the superscript * on the right-hand side of the equation. If i + 1 is substituted for * then the second derivative of the temperature with respect to distance is evaluated at the end of the next time step (using tem¬peratures not already known). Such an approach leads to a set of equations involving unknown temperatures and a solution procedure which is referred to as being implicit. The important characteristic of the implicit procedure is that it leads to a set of equations that must be solved for each time at which the temperature dis¬tribution is required. If instead of substituting (i + 1) for the superscript *, i is substituted, the following equation is obtained: [ ] {In this case the new temperature is defined in terms of an already known temperature distribution. The solution procedure is now known as explicit and has the impor¬tant characteristic that there are no equations that have to be stored or solved. A practical advantage of this

Jan 1, 1982

By Nelson R. Shaffer

Biotechnology has become a household word of the nineties, and it is expected to become as important in the next century as the computer is in the present. Numerous books and articles portray biotechnology developments as nothing less than a scientific revolution. Almost everywhere one looks new biotechnical breakthroughs are being reported that offer almost limitless opportunities to harness the force of living things to produce materials and manipulate their properties. Biotechnology has been broadly defined as any applications of biological organisms, systems, or processes to manufacturing and service industries. This seemingly new technology is, in fact, one of mankind's oldest scientific activities (Table l), which has been recently revolutionized by techniques of genetic engineering that arose out of basic research in biology, biochemistry, genetics, and information sciences. From fields as old as agriculture and medicine to those as new as monoclonal anti-bodies, transgenic plants, or biocomputers are encompassed by biotechnology. Like most human endeavors, industrial minerals play critical roles in biotechnology. In addition biotechnology holds real potential to improve extraction and beneficiation of certain industrial minerals themselves. BIOTECHNOLOGY OVERVIEW Companies using established biotechnical techniques make up large and diverse groups such as agriculture, chemicals, and pharmaceuticals. The massive US Pharmacopia (Anon., 1990a) provides detailed specifications for minerals used in medicines. Alumina, zirconia, apatite, and bioactive glass have seen service as implant materials (Williams, 1990) and new uses for minerals in health sciences are being actively researched. Agriculture produced $361 billion worth of food and drink during 1991 in the United States; organic chemicals, pharmaceuticals, and enzymes accounted for $68, $59, and $42 billion, respectively (Anon., 1992a). It is not possible to separate the contributions of industrial minerals to biotechnical products, but they represent a very large and rapidly growing new field of uses. The new biotechnology has nearly 300 small companies, plus 15 established companies with 742 biotechnology-related firms (Dibner, 1991b). Revenues exceeded $2 billion in 1990 and are expected to grow to $50 billion by 2000 (Anon., 1992c), with worldwide sales exceeding $100 billion (Burrill and Roberts, 1992). Federal research amounted to $3.4 billion in 1990 (Anon., 1992b). The United States is the world leader in biotechnology, but other countries have large, well-funded programs. Despite debate about safety, obstacles to new biotechnology products are declining (Embers, 1992, Gibbons, 1991). Fifteen biotechnology drugs valued at $1.2 billion (Thayer, 1991a) are on the market, and more than 100 are in various stages of testing (Edington, 1992). Many diagnostic tests are also in use or development (Demain, 1983, Anon., 1992). Large scale efforts to produce or transform important chemicals are also underway (Ng et al., 1983, Hinman, 1991), and research into geologic uses of biotechnology has begun. Much has been published about microbial mining, oil recovery, desulfurization, bioremediation, and other geologic aspects of biotechnology, but this chapter is the first attempt to explore interactions of biotechnology and industrial minerals. This chapter examines uses of minerals in biotechnology; how biotechnology can be used to discover, recover, and beneficiate industrial minerals; and speculates on some potential, but as yet untried uses. Definitions What exactly does the word biotechnology mean? Bud (1989) states that the first use of the term was by Karl Ereky in 1919 to cover the interaction of biology and technology, and in 1933 the term was used in Nature. After citing seven different definitions, Smith (1988) concludes that biotechnology is a series of enabling technologies involving practical applications of organisms or their subcellular components to manufacturing and service industries or to environmental management. Walker and Cox (1988) suggest a definition of "the practical applications of biological systems to the manufacturing and service industries and to the management of the environment." Primrose (1991) says that it is "the commercial exploitation of living organisms or their components." There is essentially an older broad use of the term and a new use. The US Office of Technology Assessment (Anon., 1984) uses a broad definition that includes any technique that uses living organisms (or parts of organisms) to make or modify products, to improve plants or animals, or to develop micro-organisms for specific uses. Definitions are different, but they all have several fundamental elements that include the control, management, or manipulation of living things for commercial, industrial, or useful ends. While such a definition encompasses all of agriculture in practice the "new" biotechnology is restricted to processes involving microorganisms-plant and animal cells, or enzymes. Many consider biotechnology to be recent, but it is one of our oldest technologies as evidenced by the prehistoric origin of brewing, cheese-making, and other techniques. [Table 1] gives some of the important developments in the history of biotechnology. Smith (1988) breaks down historical developments of biotechnology into four phases: 1) prehistoric with no understanding of underlying processes; 2) nonsterile processes; 3) sterile processes after 1940; and 4) genetic and recombinant DNA technology deliberate design of special organisms or processes.

Jan 1, 1994

By C. A. Rowland

Introduction Ball mills are lined drums, either cylindrical in shape or modified cylinders that have either one or both ends of the shell, consisting of conical sections, that rotate about the horizontal axis. Fig. I I shows a cylindrical mill, Fig. 12 a conical ball mill, and Fig. 13 a Tricone ball mill (Hardinge tradename). Steel or iron grinding media, generally in the shape of spheres, are used to grind the ore to the specified product size. In order to obtain more contact area for grinding and to simulate the shape of worn balls, balls have been made with two concave surfaces diametrically opposite each other. Some concentra¬tors, such as Erie Mining Co., have used slugs cut from worn and broken rods to supplement the balls in ball mills and save money otherwise lost as rod scrap. Cylindrical and conical shapes have been tried instead of balls, but balls remain as the most common shape grinding media used in ball mills. Ball mills were a logical development from the earlier pebble mills that used hard natural pebbles such as flint pebbles or sized ore pebbles (obtained from the ore itself) as grinding media. In the early 1900s36 it was found that when cast iron or cast steel balls were used in place of flint or ore pebbles, the mills drew more power and gave greater production capacity. Advances in technology have resulted in the manufacture of ball mills up to 18 ft diam inside shell, drawing up to 8,000 hp. Ball mills are employed to grind ores, especially the more abrasive ores, to finer sizes than can be produced economically in other size¬reduction machines such as roll crushers, hammer mills, and impactors. Ores can be ground dry-dry grinding-or in a slurry-wet grinding-using ball mills. Dry grinding nominally refers to less than I %v moisture by weight. If the moisture content increases by several percent, dry grinding capacity is significantly reduced as shown in Table 17. The usual range of solids content in wet ball-mill slurries is from 65 to 80% by weight. Wet grinding is used to prepare the feed material for unit opera¬tions such as flotation, magnetic separation, gravity concentration, and leaching that require a slurry of liberated valuable mineral and unwanted gangue particles. Dry grinding" is employed to produce feed for agglomeration, pelletizing, and pyrometallurgy processes that require feed that is dry or nearly so and for finely ground industrial mineral products used in the dry state. Dry grinding is also used when minerals cannot be dewatered economically to the required moisture level or when the ground product reacts unfavorably with liquids. For example, cement clinker must be ground dry. Dry grinding requires about 30% more power than wet grinding for comparable size reduction .28 The total power required in a dry¬grinding ball-mill plant including drying may be double that required for a wet-grinding plant. Grinding-media and liner consumption in dry grinding reported as pounds of metal consumed per kilowatt-hour per ton of ore" is 10-20% of that used in wet grinding. The Wabush pellet plant, Point Noire, Que.3o reported ball consumption dropped from 6.3 lb per ton of ore ground to 2.5 lb per ton of ore ground when they converted from wet to dry grinding, and a 30% increase in power consumption. A number of comparisons made on wet and dry grinding of cement raw materials show metal consumption in dry grinding to be 10% of that in wet grinding. The capital costs for wet grinding are generally lower than for dry grinding. When thickening and filtering of the wet-ground product are required, dry grinding may have a lower capital cost. With open-circuit grinding the ball-mill discharge passes directly to the next processing step without being screened or classified and no fraction is returned to the ball mill (Fig. 14). In closed-circuit grinding the ground material, undersize, in the ball-mill discharge is removed either using a screen or a classifier with the oversize being returned to the mill for additional size reduction (Fig. 15). The over¬size material that is returned to the ball mill is called the circulating load. Open-circuit ball-mill grinding requires more power than closed¬-circuit grinding for products containing similar amounts of top-size material. The less the amount of oversize allowed in the product, the longer the ore must remain in the ball mill when grinding in open circuit. This increases the production of extreme fines and thus the consumption of more power. The power required for open-circuit ball-mill grinding can be estimated using the multipliers listed in Table 18 and knowing the power required for closed-circuit grinding to yield the desired product particle size. For example, assuming the desired grind size is 90% passing some specific top size, open-¬circuit grinding would require 1.40 times the power to achieve similar results as closed-circuit grinding.

Jan 1, 1985

By R. M. White

Abstract. Gulf Coast Salt Domes are being considered for radioactive waste disposal and this paper describes the geophysical techniques currently being used and those being considered for later phases of the program. Detailed gravity work utilizing a modification of the Talwani-Ewing modeling technique, shallow high resolution reflection seismic surveys using the mini-sosie concept, surface electrical resistivity and down-hole geophysical logging to add a third dimension to all of the surface techniques and to allow inverse modeling of the surface data sets are being utilized. The logic of the combined geophysical surveys will be discussed as well as their relative economics and their relationship to the overall exploration effort.

Jan 1, 1980

In 1981, North American Degerstrom Contracting began gold mining on the Little Rocky Mountain Range, near Zortman, MT. In the past four years, the price of gold has fluctuated from a high of $16/g ($500 per oz) to a low of $9.60/g ($300 per oz). With that kind of a moving target, profitability presents a management challenge. Based in Spokane, WA, Degerstrom employs 85 people at the mine. It is one of the largest heap leach operations in the nation. Last year, 8 Mt (9 million st) of low-grade ore was moved to the leach pads in a two-shift, five-day operation. The current pit being worked produces 218 kt/m (240,000 stpm) of leaching ore. Ore fragmentation critical The gold bearing ore has no defined patterns or seams. It is an igneous intrusion resulting from volcanic activity. It occurs spontaneously throughout the range. To recover the gold, areas are staked out and core samples are taken and analyzed. These data determine if the ore is rich enough to mine. Once mining begins, an average of 150 holes are drilled and blasted on a two-day rotation. Five Ingersoll-Rand T4s drill on a 4- x 4.5-m (14- x 15-ft) spacing to a 7.6-m (25-ft) depth. "When we first came to this site, we drilled on a 3.6- x 3.6-m (12- x 12-ft) pattern," said Paul Baker, superintendent for N.A. Degerstrom Contracting. "We did some experimenting using a high density emulsion in the bottom of the hole and filling the rest with regular Anfo." The high density emulsion saved the company on overall drilling and blasting costs since the holes could be spaced on a wider pattern and still achieve the high degree of fragmentation needed for leaching. Blasting is important because the gold lays in the natural strata of the rock. Also, the ore is not crushed before it goes to the leach pads. So complete fragmentation is critical. Terrain dictates loading/hauling system The shot ore is worked in 6 m (20 ft) benches by two Caterpillar 245 front shovels equipped with 3 m3 (4 cu yd) buckets, a 992C wheel loader with a 12.6-m3 (16.5-cu yd) bucket, and a 988E high lift fed by a D9H. A Caterpillar D10 tractor trap dozes to the loader. According to Baker, the 6 m (20 ft) bench is an efficient lift for the front shovel and the wheel loader. In tight quarters or in a pocket of ore, Degerstrom uses the 245s. For high production areas, the wheel loader is used. The hauling fleet includes 28 Cat 773 off highway trucks. The 45 t (50 st) trucks are seven pass loaded by the 245s with a 25-second cycle time per bucket load. The 992C loads the trucks in two passes in less than 30 seconds. The cycle time is four to five minutes for a 1.2-km (4000-ft) haul on grades that average 10%. Grades, in some places, exceed 16%. "We use the 773s because they have a low weight-to-horsepower ratio - that's what you need in steep country," Baker said. The trucks haul the shot ore to the leach pad. Currently, Degerstrom is working a pad that has a capacity for 5 Mt (5.5 million st) of ore. Building the leach pads The base of the pads resemble a large drainage basin. They take about one month to construct. A 0.3-m (1-ft) layer of impervious bentonite clay is hauled in and leveled by a Cat D9H tractor. A 30-mm (1.2-in.) PVC liner is then laid in place on top of the clay base. It, in turn, is covered with tailings from an old on-site mill. Degerstrom uses the tailings to protect the liner from tears when the ore is dumped. The ore is leveled and built up in 9 m (30 ft) lifts by the D9H. The tractor rips the top layer. Then, a network of plastic irrigation pipe is put in place to distribute a cyanide solution over the surface. The leaching solution percolates through the ore and dissolves the gold. The solution drains from the pads and is pumped to a 22.7-ML (6-million gal) pregnant solution holding pond. The liquid then goes through two separate filtration units. One unit removes entrained solids and the other side adds a zinc dust to precipitate the gold. This is collected on filters. The affluent then returns to a solution holding pond for redistribution through the pipeline network. Typically, 80% of the pad's potential recovery takes place in the first 30 days of leaching. The process stops when the ambient temperature falls below the cyanide's freezing level. And each pad has a leach cycle of four to five years before recovery values decline to a point where further leaching becomes uneconomical. At Zortman, it takes 9 t (10 st) of ore to produce 31 g (1 oz) of gold. It is therefore critical to keep total production costs down and efficiency high.

Jan 1, 1986

By Arnolfo A. Valdivia

Prior to 1971, there were several clinical trials to evaluate programs for early detection of lung cancer. Among these, the Philadelphia Pulmonary Neoplasm Research Project,(3) the Veterans Administration Study published by Lilienfeld,(7) and the controlled trial of the Kaiser Foundation Health Plan showed an overall five year survival rate of 8% for newly detected cases (the same as the national statistic for unscreened patients). In 1971, the National Cancer Institute initiated three randomized, controlled mortality studies using lung cancer screening of persons at high risk (male smokers over 45 years old). The studies are being conducted at the Johns Hopkins University Hospital, the Mayo Clinic, and the Memorial Sloan-Kettering Cancer Center. The studies have slightly different designs in the combination of sputum cytology and chest x-rays. At the Mayo Clinic the study group is offered screening with sputum cytology and chest x-rays every four months, whereas the control group is advised to have an x-ray and cytology every year. No reminders are sent, and it is believed that only about 20% of the control group is screened. At Johns Hopkins and Memorial, both experimental and control groups are offered annual chest x-rays. The experimental group is additionally offered sputum cytology every four months.(5) At present all of the programs show that screening can detect cancers that are undetectible by other means. However, at this time mortality rates in the control and experimental groups are not significantly different in any of the three studies. OUR PROGRAM Our clinic is located in Grants, New Mexico and we provide most of the pre-employment physical examinations for the mines operating in the Grants area (Kerr McGee Nuclear, Homestake Mining, United Nuclear, Western Nuclear, and Ranchers). In the examinations, we obtain the previous mining history of the worker, a chest x-ray, a sample of sputum for cytological examination, and a blood sample. We also provide routine annual physical examinations of the workers, with special interest in the detection of bronchogenic carcinoma. In the early seventies, we did not have a definite surveillance program. We did not know whether we should have a program like the one started at Memorial or like the one started at the Mayo Clinic. After long consideration, we decided to have a program that does not demand a sputum cytology and chest x-ray every four months, but that allows as many chest x-rays and sputum cytologies as needed to diagnose lung cancer as early as possible. We believe that, if a screening method for cancer is to be optimally effective, it must detect the process at stages early enough for curative therapy. We order a test depending on the age of the miner, the race, the mining history, the smoking history, the radiation exposure levels, and the results of the previous chest x-ray and sputum cytology. With the help of the computer, we have a list of all the miners who should be watched closely because of age, race, mining history, smoking history, radiation exposure, etc. Examination of the miners is performed at our clinic, where all the records are kept. The sputum is collected there but examined in Grand Junction, Colorado, by Dr. Geno Saccomanno. There are two ways to collect sputum. The best way is to collect three consecutive morning samples. For this, we need the cooperation of the miners. They have to follow these instructions and mail the bottle containing the sample to Grand Junction. "Instructions for obtaining a good cough specimen" The enclosed plastic bottle contains a preservative solution, so do not empty out the liquid in it. When you go to bed, place the plastic bottle at your bedside where it will be handy in the morning. When you first get up in the morning (before breakfast) try to cough up some "phlegm" from deep in your chest, and spit it into the liquid in the bottle. Try coughing several times. If you have difficulty coughing, try inhaling deeply the steam from a teakettle (or home-type inhalator). Keep the amount of saliva (ordinary spit) that you put into the bottle along with the cough specimen as small as possible. Do not collect the "phlegm" or mucous that comes from the back of your nose. Put the cap back on the bottle, and shake it vigorously for two minutes. If the amount of material you have coughed up is quite small, then keep the bottle at your bedside for three or four days, and each morning try to add another cough specimen. After obtaining your cough specimen, repack the bottle in the mailing container, and attach the enclosed mailing label. It does not require any postage stamps. Unfortunately, some miners "forget" to mail the sample and end up with an incomplete physical examination. To avoid this some companies, like Homestake, request that we obtain the sample in our clinic by forcing cough and expectorant with a nebulizer machine. This method does not give as good a sputum sample as the previous one, but we do get a sputum sample for every miner. The policies of different companies, in regard to annual physical examinations are different. All

Jan 1, 1981

By Aurel Goodwin

MSHA is the regulatory Agency which administers the Federal Mine Safety and Health Act. Although MSHA's principles largely derive from this Act, they do not exculsively derive from it. One might expect that our principles are contained in the regulations we have published to implement the Act. To a degree, this is true; but the regulations we have now were developed and published before the current Act became effective. For this reason, as well as others, there are some principles that we support which are not evident in our regulations, or in the Act. One of these is the ALARA principle. This principle requires that all unnecessary exposure be avoided; that is, all exposures should be kept "As Low As Reasonably Achievable." Almost all professionals in radiation control can agree with this broad principle; but it would be difficult to get agreement on a regulation which would translate the principle into industry practice. Rather, our regulations specify a limit to exposure for any individual. For radon daughters the current limit is four working level months per year. By having an explicit limit such as this, the ALARA principle often becomes lost and the limit becomes the goal. The principles contained in the Act are equally broad in scope. The principles apply not only for radiation protection, but also for toxic substances and harmful physical agents. The Act states that standards for such substances or agents shall most adequately assure on the basis of the best available evidence that no miner will suffer material impairment of health or functional capacity, even if such miner has regular exposure to the hazards dealt with for the period of his working life. The Act also provides that, although protection of health must be our foremost concern, other considerations shall be the latest scientific data in the field and the feasibility of the standards. Our regulations reflected similar principles when they were developed and promulgated. From experience with other health and safety laws, Congress realized that setting and meeting an exposure limit may not be sufficient to prevent disease. The Act, therefore, contains a detailed description of additional provisions to be included, where appropriate, in mandatory safety and health standards. One of these additional provisions requires that miners be informed about the nature of the hazards associated with their job and about the means for their own protection. Another provision requires the use of labels and other forms of warning to inform miners about the hazards, about proper precautions for safe use or exposure, about relevant symptoms of overexposure, and about emergency treatment. We have partially implemented these provisions through regulation. The Act requires also that standards shall prescribe protective equipment and control of technological procedures and that they shall provide for the monitoring of miners' exposures. We have implemented this provision to a degree also through regulation. Finally, the Act specifies that, where appropriate, a mandatory standard must prescribe the type and frequency of medical examinations in order to most effectively determine whether the health of miners is adversely affected by exposure. Additionally, when a determination is made that a miner may suffer material impairment of health or functional capacity by reason of exposure, that miner must be removed from such exposure and reassigned. In order to encourage miners to take the medical examination, the Act also provides that the miner shall not suffer loss of pay as a result of being reassigned. This additional provision on medical examinations has not been implemented by regulation for radiation hazards, nor for most other toxic substances or hazardous physical agents. We believe that this conference will provide us with valuable information on medical examinations for radiation exposure, as well as on other considerations to be used in future regulations. We still question whether our basic exposure standard of four working level months per year is adequate to protect a miner's health. This conference is indeed timely because both MSHA and NIOSH have been reviewing this issue for some time. I would like to reemphasize MSHA's commitment to education and training. MSHA strongly believes that miners should be informed fully about the real and potential hazards associated with their work. They should know the nature of the hazard and the means for protecting themselves from the hazard. The Act also emphasizes the need for training miners by requiring that new miners receive 40 hours of training before working underground and 24 hours of training before working on the surface. The Act also requires eight hours of refresher training annually. This training must include information on the relevant health hazards and the potential consequences of overexposure. Before I close, I would like to mention briefly another source of principles that MSHA often consults to obtain guidance for both the development and enforcement of regulations. These are the various court decisions that are handed down from time to time. Two in particular that I want to stress are the recent Supreme Court Decisions on benzene and cotton dust. Although neither of the decisions has a direct impact on mining or radiation, they do clarify some general principles and directions for regulatory agencies, such as the role of cost-benefit analysis and the need for risk assessment. In closing, I wish us all an open exchange of information for a productive and useful conference.

Jan 1, 1981

By Charles R. Nelson

INTRODUCTION The support of openings in weak sandstones can be achieved by spraying on a liquid grout which will soak in and harden to form a shell. In the St. Peter sandstone of the Minneapolis-St. Paul area of Minnesota compressive strength is increased from about 0.7 MPa (100 psi) to over 7 MPa (1000 psi) to depths of 152 mm (6 in.). A shell of this strength and thickness can pro¬vide support much as shotcrete does but at lower cost. If the spray-grouted shell is inadequate for the support needed, it provides a good base for shotcrete application, which is often difficult in weak or friable sand¬stones because the fresh shotcrete peels off along with a thin layer of sandstone. The spray-grouted technique was initially developed at the Civil and Mineral Engineering Dept. of the University of Minnesota using funds provided by RANN Div. of the National Science Foundation. It was used as temporary support in a 3-m (10-ft) wide, 3700-m (12,000-ft) long storm water drain tunnel by the Minneapolis Sewer Construction Dept. Much of the field development at this construction project was carried out with the cooperation of the personnel of the city of Minneapolis. It also was used for temporary support in a 3.3-m (11-ft) wide, 1800-m (6000-ft) long storm water tunnel built in 1978-1980 for the Minnesota Dept. of Transportation in Minneapolis. In 1975, 2000 m (6500 ft) of 1.5-m (5-ft) wide utility tunnels were spray-grouted for final lining. A 50-m (150-ft) long, 1.2-m (4-ft) wide test tunnel built at the university in 1974 was sprayed for half its length and is being observed for long-term behavior which to date (1981) is good. GROUT MATERIALS The requirements for the liquid grout are more severe than those for injection grouting. Besides being able to penetrate the rock and develop sufficient strength, it must have the following properties: (1) be nontoxic, noncombustible, and have a low odor for spraying under¬ground; (2) have controllable viscosity and setting time for adjusting to the permeability of the sandstone and the required depth of penetration; and (3) must "wet" the sandstone and develop capillary "draw" to penetrate to the required depth. A sodium silicate-based grout with the proper setting agent meets these requirements in the local St. Peter sandstone which has a uniform grain size dis¬tribution curve with D,,, = 0.1 mm, permeability of about 10 to 20 darcys, a porosity of about 25%, and less than 5% of silt size or smaller. It is 97 to 99% pure SiO2. The grout mix consists of Philadelphia Quartz Co. Type N sodium silicate and Celtite 55 Terraset Spray Grade (SG) distributed by Celtite Inc., Cleveland, OH. A volume mix ratio of 100:9:125 for sodium silicate to setting agent to water produces a grout with two to four centipoise viscosity, and initial set at 20°C of 20 min. Setting time is temperature dependent, being longer at lower temperatures, but can be adjusted by varying the setting agent concentration a few percentage points. APPLICATION The application is simple. Spray the grout on the surface until the desired penetration is achieved. Penetration rates of about 5 mm (0.2 in.) per min. are realized in the St. Peter sandstone for up to 152 mm (6 in.) of penetration (30 min. of spraying). The set time must be longer than the spraying time. The spray nozzle should be moved back and forth at a rate that minimizes runoff due to surface buildup. Both a single large hand-propelled nozzle and multiple small machine propelled nozzles have been used successfully. The pumping and mixing equipment used consisted of: (1) a hand-held, hand-pumped, 11-L (3-gal) gar¬den sprayer for test patches (mix by shaking); (2) 76-L (20 gal) batch mixing in barrels with small electric pumps for spraying and mixing; (3) two component pumps consisting of Hypro Model 5300 piston pumps (Hypro Pump Co., St. Paul, MN) on a common shaft driven by an electric motor. The sodium silicate is premixed with a portion of the water and the setting agent is mixed with the rest of the water. The two components are combined and mixed just before the nozzle; and (4) a proportioner metering the setting agent just before the nozzle with the water and sodium silicate premixed and pumped from the surface. The proportioner permits the bulk of the grout to be pumped through a single pipe or hose with only the setting agent stored underground. This is efficient and low in cost. COST The grout mix costs about $0.20/L ($0.75 per gal) in 1978 dollars. Filling all the voids 100 mm (4 in.) deep (25% initial porosity) would have a materials cost of about $5.40/m2 ($0.50 per sq ft). The labor cost of application is usually less than the material cost for normal tunnel jobs. A pump and proportioner cost less than $1000. Sodium silicate storage tanks and pip¬ing costs would depend on the particular site conditions. REFERENCES AND BIBLIOGRAPHY Nelson, C. R., 1977, "Spray Grouting for Tunnel Support and Lining," Underground Space, Vol. 1, No. 3, pp. 241¬ 245. Yardley, D. H., Nelson, C. R., Stocker, T. H., 1974, "So¬dium Silicate Spray Impregnation of Tunnels in the St. Peter Sandstone," Research Report, University of Min¬nesota, 118 pp.

Jan 1, 1982

By M. Bleeck

Calcium carbonate (CaCO3) is one of the most ubiquitous and versatile minerals found in the earth's crust. Its availability, attractive physical properties and relatively low processing cost make CaCO3 the most widely used filler material today. It is mined in three different forms - chalk, limestone and marble. Each physical form of CaCO3 has different qualities due to differences in postdepositional geology. But the chemical composition remains the same, with CaCO3 an inert component of the finished product. In the past, the paper industry largely left CaCO3 by the wayside, as it cannot withstand the acid-based papermaking process. But conversion to an alkaline system by many US mills changed this picture. Carbonate suppliers have put time and effort into research and development, demolishing barriers and creating new possibilities for what is a simple, natural product. By controlling particle size, size distribution and particle charge, the industry uses ground calcium carbonate (GCC) as a performance enhancer and as an extender for more expensive ingredients. It is estimated that the United States uses 3.6 to 4.1 Mt/a (4 to 4.5 million stpy) of CaCO3. Consolidations and mergers are taking place in the industry. Of the 12 major GCC producers in operation nine years ago, seven are left. Mineral Technology (US) is the dominant precipitated calcium carbonate (PCC) producer with more than 50 satellite plants worldwide. Other producers include Georgia Marble (French); Franklin Industries (US); OMYA (Swiss); J.M. Huber (US); ECC (English); and Filler Products (US). Global Stone PenRoc (Canada) is the only newcomer. In addition to this group, there are three small producers left in North America, each with a capacity of less than 100 kt/ a (.110,000 stpy). The trade organization operating as the Pulverized Limestone Division of the National Stone Association renamed itself the Pulverized Mineral Division, to increase its membership pool. The US paper industry is a predominant GCC con¬sumer, using approximately 800 kt/a (882,000 stpy) at an approximate cost of $130/t ($1.43/st). European paper mills pioneered alkaline papermaking. In the early 1960s, they began using GCC as filler and soon thereafter added GCC to their coating formulations. A decade later, the North American paper industry followed suit. The conversion from acid to alkaline paper production benefits the economic and performance aspects of the industry. Less pulp is needed, paper machine maintenance and effluent treatment costs are reduced, and sheet strength, opacity and brightness are increased. Perhaps most important to the reader, the sheet is desensitized to ultraviolet light, extending the paper's archival ability. CaCO3 can provide the papermaker with additional control of his paper. For example, PCC has long supplied the tobacco industry with a means to slow down the burning rate of cigarettes. Due to enhanced performance with regard to bulk and opacity, filler PCC use has risen to 1,500 kt/a (1,650 stpy) in the United States, at an approximate price of $130/t ($143/st). The majority is produced onsite at the paper mill, using "satellite plants." This concept reduces freight cost because only quicklime (CaO) is shipped to the mill, not CaCO3 slurry. The future of CaCO3 is encouraging. The amount of natural ground CaCO3 used is expected to double by the year 2005 to approximately 8 Mt (8.8 million st) worldwide. Acid papermaking practices will feel an increasing pressure to convert to an alkaline process as larger volumes of GCC containing paper enter the recycling market. CaCO3 reserves are plentiful. They will supply the ever growing demand for increasingly sophisticated paper. The plastics industry is supplied with almost 900 kt/ a (990,000 stpy) GCC at an annual growth rate of 4% to 5%. The price of a functional, inorganic filler, surface modified for the plastics industry, has an average selling price of $220/t ($243/st). GCC represents the most common filler, creating a product with higher gloss, better dielectric properties, impact resistance, weatherability and shrinkage control. CaCO3-filled plastics surround us - auto hubcaps and dashboards, shower enclosures, floor tiles, wire coatings, microwave dishes and Tupperware. The caulking and sealant industry is an enormous GCC user, with annual consumption requiring 1.13 Mt (1.2 million st) at about $44/t ($48.50/ st). Caulking and sealant may be highly filled with GCC yet undergo no adverse flow effects, with a narrow particle-size-distribution filler decreasing the binder demand. The CaCO3 industry, as well as the carpet industry, are more or less tied to the growth rates of the construction industry. It is estimated that the carpet industry uses some 680 kt/a (750,000 stpy) of GCC at about $25/t ($27.50/st). The paint industry uses approximately 300 kt/a (331,000 stpy) at a 1 % to 1.5% annual growth rate. Here, too, GCC is the dominant filler. It is used to enhance flow characteristics and color uniformity. It also extends costly titanium dioxide, creates sheen and controls roughness, hardness and tack. More CaCO3 should be used in the future as industry shifts from solvent-free or water-based formulations that can accommodate higher GCC volumes. CaCO3 is not imported or exported in any great quantity. Most areas have reserves of their own and the selling price is relatively low. Even so, quality varies from deposit to deposit. As our needs become morespecific, it remains a challenge to provide varying industries with a fitting product.

Jan 1, 1998

By Bernard F. Anderson

INTRODUCTION In this chapter, the term "down-the-hole drill" (DTH drill) is used as a generic name that encompasses the various trade names and other references such as "downhole drill," "in-the-hole drill," etc. This chapter is limited to a description of DTH drills used in stoping large underground ore bodies. DTH drills differ from conventional drills by virtue of the placement of the drill in the drill string. The DTH drill follows immediately behind the bit into the hole, rather than remaining on the feed as with ordinary drifters. Thus, no energy is dissipated through the steel or couplings, and the penetration rate is nearly constant, regardless of the depth of the hole. Since the drill must operate on compressed air and tolerates only small amounts of water, cuttings are flushed either by air with water-mist injection or by standard mine air with a dust collector at the collar. HISTORICAL DEVELOPMENT Mine managers have long known the economies enjoyed by quarry and open-pit operators in producing large quantities of ore. The savings are due primarily to the availability of massive equipment, capable of drilling large blastholes to reduce the amount of drilling, increase the fragmentation, reduce secondary blasting, and im¬prove the flow of the product. In an attempt to reduce underground mining costs, various methods are used for long-hole drilling, includ¬ing standard pneumatic percussion drifters and diamond drills. These systems have their shortcomings; percus¬sion drills are limited to small hole sizes and they ex¬perience excessive deviation and significant loss of energy with increased depth. The diamond drills provide deeper and straighter holes, but only at high cost. Both systems suffer from high noise levels, low penetration rates, and poor explosives distribution, among other problems. When the mining companies approached the drill manufacturers for a compact and portable large-hole jumbo for underground use, they specified not more than 1 % deviation on 60 m (200 ft) of vertical hole and a penetration rate of 15 m/h (50 fph). On Dec. 23, 1960, a test unit was placed in service in Montana and met the performance criteria. Though lacking the so¬phisticated features available today, the economies of surface blasting were brought underground. Unfortunately, the first system did not gain immedi¬ate acceptance in the industry. Among the factors con¬tributing to its demise were resistance to change, the need to alter development methods for the ore bodies, and a lack of flexibility in moving the rig from setup to setup and from level to level. In 1972, the mining industry again challenged the drill manufacturers to provide a workable jumbo that would combine compactness, ease of maintenance, relia¬bility, and efficiency, all on a self-propelled chassis. The manufacturers responded by providing improved jumbos, which have been accepted with enthusiasm throughout the mining industry. Today's DTH jumbos are capable of drilling from 100 to 200 mm (4 to 8 in.) diam holes that can be reamed to even larger diameters. The holes can be drilled to depths of 150 m (500 ft), depending upon ground conditions and the capability of the jumbo to retrieve the steel and drill. Fig. 1 illustrates a typical DTH jumbo. APPLICATIONS The uses to which DTH drill jumbos have been put are quite numerous, with new uses being found regularly. For convenience, these uses may be classified as primary blastholes and nonblasting holes. Primary Blastholes The original purpose for the development of the DTH jumbo was for drilling primary blastholes that could be mined by open-stope methods. Prior to the advent of the DTH jumbo, extensive development was required before production drilling could begin. Sub¬levels were required to allow access for column-and-arm stopers or ring/fan jumbos, to the extent necessary based on the effective penetration of the chosen machine. With the DTH jumbo, the mine engineer is able to reduce preproduction time and development costs. How¬ever, the most significant saving results from an im¬proved cost per ton of broken ore in the production phase. To utilize a DTH system, only a top heading and drawpoints are necessary. The top heading can be the width of the ore body with a 3.7-m (12-ft) back. A drop-raise pattern is drilled and shot to begin the stoping operation, providing a free face for subsequent blasting. A typical layout is illustrated in Fig. 2. The advantages of this system include: 1) Drilling and blasting are independent operations, and blasting can be performed at a rate congruous with the mine's ton-per-day capacity. 2) The development layout is simplified. 3) Good explosive distribution is achieved, provid¬ing more uniform fragmentation. 4) Environmental conditions for operators are im¬proved, including improved safety with all work directed downward (not overhead), lower noise levels, little fog, and a reduced dust count. 5) Improved production per manshift. 6) Simplified and easier operator work cycles. 7) Reduced cost per ton of product. 8) Fewer holes lost due to ground shifts. Nonblasting Holes With the introduction of the compact DTH jumbos, other practical uses became apparent, including the drilling of: 1) Holes for sand fill, from level to level and from level to stope. 2) Drain and dewatering holes. 3) Power and communications cable holes.

Jan 1, 1982