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By R. W. Henn

Introduction Shotcrete, by the American Concrete Institute definition, is mortar or concrete pneumatically projected at high velocity onto a surface. The term gunite is still common in literature. It is a trade name established by Cement Gun Co. of Allentown, PA, around 1911 for a dry mix of fine aggregate material, hydrated at the nozzle, and pneumatically projected at high velocity onto a surface. This article will discuss shotcrete applications, ingredients, mix proportioning, batching and mixing, admixtures, testing, and steel fiber vs. conventional shotcrete. Applications Shotcrete is used primarily by the mining and underground construction industries to temporarily stabilize an excavated rock surface until a permanent cover can be installed. It may also be used as the permanent covering material. Other uses include refractory material, architectural finishes, slope stability, wearing surfaces, repairing existing concrete structures, fireproofing, and corrosion protection of steel. Wet Versus Dry Process In the wet process, all ingredients, including water but not accelerator, are mixed the same way as normal concrete before its placement by pneumatic projection. The accelerator is added in a liquid state through the nozzle at the point of placement. In the dry process, all ingredients, including acceleration in dry form but not water, are mixed just before placement by pneumatic projection. Water is added through the nozzle at the point of placement. Liquid accelerator may be added at the nozzle in the dry process. Ingredients Shotcrete consists basically of the same ingredients as concrete with some limitations. Type I, II, and III cements are used. However, Type II seems to be used most in underground applications. This can be attributed to ground water at most sites having slight to moderate concentrations of sulfates. Air-entraining cement or air-entraining admixtures are generally not used in dry mix shotcrete. It is generally agreed that it is impossible to entrain air due to the absence of the usual concrete mixing action, and the high velocity of impingement of the material on the application surface. Air-entraining cement or air-entraining admixtures are used in wet mix shotcrete. Their use, however, is not widespread due to the placement method and the fact that high velocity air drives off most of the entrained air. Test results are not conclusive in showing any durability increase of shotcrete to freezing and thawing by using entrained air. In spite of current practice, the trend is toward additional research, with more agencies and owners requiring the use of entrained air. The philosophy behind this is that some entrained air is better than no entrained air to increase durability to freezing and thawing. The aggregates used in shotcrete are the same as those used in concrete. However, the largest coarse aggregate presently used is 19 mm (0.75 in.). A typical gradation requirement is shown in Table 1. The water quality requirement for shotcrete is the same as that for concrete. Accelerators are by far the most important admixture in shotcrete.

Jan 6, 1984

By S. Chander

The effect of sample storage in argon and in air on kinetics of oxidation of pyrite from a coal source was studied using accelerated weathering test in laboratory columns. Prior to the weathering tests the samples were crushed, ground, and stored for various time periods in argon and air. The results showed a dramatic difference in the rate of oxidation of the aged samples. The differences were interpreted based on the hypothesis that the aging of pyrite led to the formation of surface coatings which changed with time. The significance of these results in pyrite/reagent interactions in flotation and in evaluating the acid mine drainage potential of pyritic waste materials is discussed.

Jan 1, 1993

By Antonio Peralta

Mining activity in Mexico began in the 16th century when the Spanish realized the mining potential of the new territory. Since that time, mining has evolved into a strong and vital industry for the country. Good professionals in the mining-related fields are fundamental to the growth and improvement of this industry. Since the end of the 18th century, formal education in mining has been available at the University of Guanajuato, in Guanajuato, Mexico. This school created one of the first schools of mines on the North American continent in 1828. In recent times, some other programs have been created at other universities. Universities offering mining related programs Listed here are the seven universities in Mexico that offer mining related degrees. They are the University of Guanajuato, UNAM, the University of Zacatecas, the University of Chihuahua, the University of Sonora, the University of Coahuila and the University of San Luis Potosi. The University of Guanajuato is the only Mexican university that offers graduate studies at the master's level. No PhD programs in mining engineering are offered at any Mexican university. Programs at each university are of different duration. They are: • University of Guanajuato-Five-and-a-half years, divided into five years of classroom instruction and six months in the working place. • University of Zacatecas - Five years of classroom instruction.

Jan 1, 1994

By Walter G. Steblez

The countries of Eastern Europe are former members of the Council for Mutual Economic Assistance (CMEA). Geographically, they are located in Central Europe (Poland, Hungary, Czechoslovakia) and Southeastern Europe, largely in the Balkans (Yugoslavia*, Bulgaria, Romania and Albania*). It is important to remember that all of the countries of Eastern Europe had central planning imposed on them by the Soviet Union during the years following World War II. "Stalinist" policy for industrial development had three important tenets: • central economic planning, or a nonmarket-based command economy with administratively set prices; • economic priorities that stressed the development of heavy industries such as mining, metallurgy (steel) and heavy machine building; and • self-sufficiency at all costs, which was aimed at abolishing reliance on the world capitalist market, or reducing it to a minimum. Essentially, the system of industrialization that was used in the Soviet Union in the 1930s and 1940s was imposed on Eastern Europe. Organic market economic development in Eastern Europe ceased during this period. The flaws of this experiment became apparent when mineral processing facilities were set up in many areas of Eastern Europe that did not have reliable resources. Poland, Hungary and Romania are good examples of this situation. The Soviet Union became the principal supplier of iron ore and other minerals and mineral fuels, not only to these countries but to the Bloc as a whole.

Jan 1, 1992

By Yutaka Mochida

OUTLINE Beginning in 1946, a careful geological survey was made of the Tsugaru Straits. In 1964, excavation of the exploratory tunnels were started to observe the characteristics of faults and rock formation directly, and to study construction methods of excavation and sealing water inflow. In March 1971, the investigation stages were over, and in April 1971, the construction project was authorized from the government. The geological features in the straits may be divided into three broad classes. One third consists of andesite on Honshu side. The other two thirds consists of tuff. The middle of the straits consists of tuff, so-called Kuromatsunai-layers. On Hokkaido side, it consists of tuff, called as Kunnui-layers. There are nine noteworthy faults on the route. Prior to the excavation of the main tunnel, the horizontal pilot tunnels and service tunnels are made. Exploratory borings are taken to confirm the geological conditions and the presence of gushing water. Groutings for sealing water inflow are carried out when necessary. In geologically bad conditioned sections such as disintegrated or fractured zones, special methods are applied. In geologically good zones, the tunnels are bored through using the same methods as for a mountain tunnel driving. The pilot tunnel being constructed at the sea bottom is under direct control of the Japan Railway Construction Public Corporation. The service tunnel and the main tunnel are being excavated by contractors.

Jan 1, 1981

By Takeshi Tanaka, Shinya Ono, Tsuyoshi Kato, Masaji Kato, Yuusaku Tominaga

The knowledge of the macroscopic characteristics of a ventilation network is of fundamental importance for taking counter- measures to increase air quantity in-specified branches, such as those corresponding to working faces at advancing entries or longwall panels. Because the ventilation characteristics are inherent properties of each mine, it takes a long time for an engineer to accumulate experience on how a specific regulator may affect the air quantity in a specific branch of a mine with a complicated ventilation system. A computer-simulation training system, based on the macroscopic ventilation characteristics, is presented in this paper. When a trainee selects a specific condition from five regulator option (closed, almost closed, slightly open, half open, open), the change in air quantity for a network branch is indicated by the color allocated to the ratio of the new air quantity to the old or planned air volume. After the ventilation simulation, the colors in the specified branches are defined by using the fuzzy set theory. The trainee continues practicing the selection of closed/open conditions for different regulators until each of the specified branches is displayed in good state of airflow. The ventilation net- work of the Horonai coal mine is used as an example of the training system to control air quantities in six working faces with 33 regulators.

Jan 1, 1996

By Sandy Clamage

"Seeing the elephant" was a term often heard from 1849 to 1859, the period of the great California Gold Rush. The phrase was used by those who dreamed of instant wealth that was believed could be found in California's virgin earth and mighty rivers. When news of fabulous gold strikes throughout the Mother Lode reached far away places, thousands of gold seekers left home and family to join in this incredible adventure. The barren, placid state of California soon became one of the most bustling, populous areas in the West. Before the Rush Before 1848, Mexico ruled California. It was then a remote region, isolated from the urbane world of the East by 2900 km (1800 miles) of scorching desert and impassible mountains. By sea, it was 29 000 km (18,000 miles) via Cape Horn. Spaniards and Mexicans settled the coastal lands on tracts donated to them by the Mexican government. The inland area was the domain of Indians and a few settlers who also had been granted estates by the government. Among those settlers was Captain John Sutter. He had been fortunate enough to obtain about 11 leagues of land (spanning 100 km or 60 miles in length) from Governor Alvarado of Mexico. Sutter was somewhat of a visionary and possessed a gifted tongue. He managed to receive most of what he desired with minimal trouble. Enchanted with Sutter's personality, Alvarado promised him Mexican citizenship. Alvarado also commissioned him to become "representative of the government and officer of justice on the Northern frontier." With uncanny foresight, Sutter began building a fortress on his land in the Sacramento Valley. It later became known as Sutter's Fort. It was more than just a fort, however. It also became a stopping place and home for many foreigners and Americans who traveled to California, and it eventually became the mecca for overland pilgrims. Sutter engaged in many projects. These included a cattle ranch, fur trade and trapping, and the construction of an irrigation system. With the influx of immigrants over the years, the community around the fort grew rapidly. Lumber was much in demand and Captain Sutter agreed to supply the necessary capital for a community sawmill. The site chosen was some 72 km (45 miles) northeast of the fort on the south fork of the American River in the town of Coloma. During the course of its construction, Sutter's carpenter, James Marshall, discovered a problem with the tailrace on the mill. First Strike Early in the morning on Jan. 24, 1848, Marshall was checking the tailrace. He noticed an object shining in a foot of water at the bottom of a ditch. He reached down and picked it up. His heart began to thump wildly as he realized he was holding a gold nugget. Sutter planned to keep the find a secret for as long as possible. He knew very well that he could become an extremely wealthy man. The ever loyal Marshall remained faithful to his employer and the pair did their best to obtain more

Jan 3, 1985

By Peter J. Huck, Bruce G. Stegman, Richard E. Thill

Introduction Backfilling of mine voids is used to prevent or control the effects of subsidence on surface structures. The pumped-slurry, blind backfilling process has been successful for stabilizing ground over abandoned and inaccessible underground coal mines. This article describes the systematic evaluation of potential monitoring systems that would improve backfill monitoring technology. It also details field proof-of-concept tests that were conducted on two systems and the difficulties encountered in implementing them. Ground subsidence over abandoned mines can have devastating effects on overlying urban areas decades after the closing of room-and-pillar mines. The result may be cracked foundations, ruptured gas and water lines, broken sewers, distortion or cracking in superstructures, and sinks and potholes in the ground surface. About 32,000 km2 (8 million acres) of land in the US has been undermined for coal. According to the US General Accounting Office, 8,100 km2 (2 million acres) have undergone subsidence and another 8,100 km2 (2 million acres) are expected to subside by the year 2000. Actual figures on land affected or threatened by subsidence are probably much higher, since these estimates were based on late 1960s studies. For more than two decades, the US Bureau of Mines has been engaged in stabilizing ground in urban areas having high risk potential for subsidence costing millions of dollars. Thousands of acres have been stabilized, protecting property worth hundreds of millions of dollars. In earlier years, backfilling was by in-mine stowing in accessible mines or by gravity-feed sluicing methods in inaccessible mines. Pumped-slurry methods have been used extensively in the past 10 years. More slurry can be injected for fill from fewer injection holes over extensive areas in abandoned mines. Since this method requires fewer boreholes, there is less disruption of surface facilities. An investigation was conducted to determine the feasibility of using advanced, remote sensing or monitoring technology for application in monitoring fill placement. Specifically, the objective was to develop conceptual systems for blind backfill monitoring that require fewer boreholes and give better definition than is now possible. Blind Backfilling Methods Blind backfilling operations are conducted from the surface and do not require personnel and equipment underground. These methods are applicable to abandoned, inaccessible workings. Two categories of blind back-filling are point support, practiced by civil engineers to protect individual structures or surface facilities, and areal backfilling methods, used to protect large areas against subsidence. Point Support Methods Point support methods are usually gravity-feed systems. In general, point support methods use small volumes of expensive fill material and require a large number of boreholes within a constructed site. Because of the processed material used for back-filling and the close spacing of the injection boreholes, control and monitoring often may be accomplished by borehole cameras. Areal Backfilling Areal backfilling is conducted mainly by the pumped-slurry injection process (Fig. 1). In eastern and interior coalfields, material used is often mine refuse or flyash, both being undesirable on the surface and costly to dispose of in an environmentally acceptable manner. Hence, the use of

Jan 12, 1983

By Robert C. Sutter

In 1958, it became obvious to some of us in the solution mining business that there were certain problems in connection with our sodium chloride brining operations which were industry- wide. These were the types of problems which, while they were important to the safe and economical operation of our salt brining operations, were not such that their solutions would give an individual company a great competitive advantage over his competitor; and since they involve the safe and efficient mining of a natural resource, it was obvious that the answers would, in any event, be shared industry-wide. Since we were dealing with underground solution mining and, consequently, unseen and remotely controlled operations, the problems are formidable and the solutions difficult. While all of us were doing work to improve our operations, none of us could justify the expenditures necessary for basic research in these areas. Consequently, the idea was conceived of research sponsored by many companies with the same problems. In 1960, 11 companies entered into a' formal agreement to jointly support research on a specific problem of sodium chloride solution mining. The project to be undertaken was the development of a method for measuring the size and extent of underground solution cavities in sedimentary salt beds. It was agreed that all research was to be performed by independent research organizations. A part-time consultant or administrator was retained to administer the research contracts and to provide technical advice.

Jan 1, 1966

By Michael R. Birkos

A new electrotelluric prospecting technique allows the mapping of ore body distribution, lithologic contacts, and voids with accurracy comparable to core drilling. Data is available almost immediately and presented in a format similar to a downhole log. This is done with a passive surface receiver that transduces electrotelluric radiation into an audible signal. Changes in the character of the audible signal lends insight into lithology, porosity, fluid type and medium. Three case histories are presented. The first delineated a paleochannel within a coal seam. The second traced old mine workings and backfill areas. The third identified a high-pressure karst void.

Jan 1, 1986

By Wallace W. Roepke, James C. Church

Introduction The history of mining machine development has been and is one of "bigger is better." While this may not always be true, it is currently the only approach open to machine designers without specific in situ cutting force information. The US Bureau of Mines has tried to modify this approach by supporting research, both in the laboratory and in field tests, that demonstrates that bigger is only better if the equipment is used correctly. In-house research has shown conclusively that deeper cutting reduces both dust and energy (Roepke, Lindroth, and Myren, 1976). This approach has been applied to a large Lee-Norse (LN456) continuous mining machine, which confirmed that slow deep cutting is beneficial and does not have a negative effect on production (Black and Rounds, 1977; Black et al., 1978). Specific in situ engineering data needed on coal cutting forces have not, however, been obtained by field tests with the LN456. Although the machine was instrumented to obtain data on one seam, it would be impractical to move the machine from mine to mine. It became apparent that a portable, easy to use device was needed for obtaining coal cutting forces in situ. Such a device could provide the engineering data needed by manufacturers, the site-specific cutting information needed by operators, and a field research tool to establish correlation with laboratory cutting tests. Design and Development According to a set of design guidelines established by the Bureau, the proposed device should be able to measure cutting forces directly with standard coal cutting bits, be transportable by two miners, and be self-contained. These guidelines were given to Ingersoll-Rand Research Inc. (IRRI). The contractor performed a rigorous review of 16 existing in situ portable tester designs, none of which could use a standard coal cutting bit to measure cutting force directly. These included a wide range of devices, all of which are discussed in the final report (Banerjee and Wisner, 1982). A series of design tradeoffs were established to satisfy initial design guidelines. The final deci¬sion was heavily weighted by an industry survey to determine the preferred concept from four man¬ufacturers and four operators. To assure ease of use with maximum information, all respondents wanted a linear cutting device that could work in seams down to 1 m (40 in.). The final design choice was between a linear and a rotary tester. The rotary tester would cut on an arc in the same manner as a rotary drum machine. It had several faults, however, that could not be resolved within the scope of this contract. The worst was related to arc cutting (Roepke, Lindroth, and Myren, 1976). Since any bit cutting on arc enters the coal at zero depth, goes to maximum depth only at the horizontal centerline, and exits at zero, maximum cutting force is obtained only at the instant the bit crosses the horizontal centerline. This centerline peak response would require that many replications be made with an arc-type cutter to get reliable force information. The arc-type tester would require extensive face preparation to establish the cutting path and extensive wasted clearing cuts to advance the device for additional layers of cuts. The linear tester could be mounted directly on any face or rib with only minor preparation. By its nature, a linear cut at a constant depth provides multiple peak force data over the length of the cut, thus reducing the required number of replications. The final review of these two major design concepts plus indus-

Jan 9, 1983

By Wallace W. Roepke, Richard C. Olson

To adequately attack the longwall dust problem, the US Bureau of Mines funded a contract to define dust sources on a longwall mine face area. The study examined the shearer during sumping, cutting, and tramming, the conveyor, and the roof supports. Tests showed the shearer to be the major dust source. Introduction Long-term illnesses place a heavy burden on miners, the mining industry, and taxpayers. Black lung payments to more than 400,000 miners and their survivors exceeded $9 billion between 1970 and 1980. They are continuing at a rate of $1 billion per year. Recent MSHA data indicate more than 50% of the longwall faces are still out of compliance with the 2 mg/m3 (0.00007 oz per cu ft) respirable dust regulation. Allowable dust limits are often further reduced owing to the presence of silica dust. This substantially increases the problem's severity. Research in the 1970s showed that by modifying the mining method through changing drum-type mining machines from high rpm shallow cutting to slow rpm deep cutting, both specific dust and energy could be reduced (Black et al., 1978; Ludlow and Wilson, 1982; Roepke, Lindroth, and Myren, 1976). The coal mining industry is using more longwall equipment to economically obtain high productivity. This high production rate, however, makes it difficult for longwalls to meet the legal dust standard. Research is showing how both these concerns may be addressed at the mining face. Results reported here explain the relationship of various major elements of the longwall system to dust generation on the longwall face. This paper is based on work completed by Foster-Miller Associates Inc. Objectives were to differentiate dust sources on the longwall face and to quantify dust levels for various elements of the shearer drum. The work was accomplished in two phases. Jonathon Ludlow, project engineer for FMA, developed and performed the necessary testing (Ludlow and Marshall, 1983). Bradley, Hadden, and Weston (1983) have done evaluations on British coal mines that produced similar results to those reported here. They found causative rela¬tionships between haulage speed, production, depth cut, drum speed, and dust generation. They also found that roof support movement increased the dustiness, whereas this work indicated that support movement had only a minor effect on total dust exposure. Both of these results are valid when the variation of mining conditions is analyzed. British mines are deeper than US mines. Therefore, roof rocks and coal are more fragmented than in the shallower US mines. Other reasons for the dust differences include the weaker shales found in Britain and the variation of roof support equipment designs between the UK and the US. Four mines in this report had rock roofs and, thus, would be less dusty. The fifth longwall mine had a coal roof, but the large cross-sectional area and high air velocity minimized dust exposure. Mines and Equipment Five underground longwall mine sites were tested (Fig. 1). All were using double-ended ranging-drum shearers (derds) with some variations in equipment. The mines and equipment details are shown in Table 1. No modification of mining equipment or operating cycles was made during the first phase of testing. The mining process observed followed the normal mine operating procedures. Three GCA RAM-1 respirable dust monitors and two vortex anemometers with cycle notations on portable tape recorders were used in each face area. The start and end of each cycle segment were recorded. Dust monitors record the dust levels once every 10 seconds, thus providing a relatively continuous real time indication of change in

Jan 8, 1984

By H. Para, A. Aguirre, G. Artadi

Mining is a very important economical activity in Peru which has been recently developed without considering environmental aspects. Here, we show an experience at the Milpo mine, situated 4,200 meters above sea level; a mining company that has the value of being one of few from the last decade that is engaged in consolidating a model of environmental protection. The model includes aspects of contamination prevention and control, with bioindicators such as pasture cultivation, vegetables, forests, and hydrobiological species breeding. Socially, through means of development programs in natural areas, forestation and support to rural agriculture, together with environmental education, it is being demonstrated that it is possible to ensure tangible results in environmental conservation and the well-being of the populations around the mine.

Jan 1, 1998

By Nicholas Kyriazia, John P. Shubilla

A performance study of oxygen self-rescuers of varying durations from the United States and other countries was undertaken as an assessment of present worldwide technology. The apparatus were tested on a Breathing and Metabolic Simulator in the Life Support laboratories of the Bureau's Pitts- burgh Research Center. Parameters monitored during the testing were inhaled levels of C02, 02, inhaled gas temperature, and breathing resistance. The metabolic demand placed on the apparatus represented the average demand of the 50th percentile miner performing a 60-min man-test 4, as described in 30 CFR 11H. Results presented include apparatus service life and subjective comments as well as averages and peaks of monitored parameters.

Jan 1, 1986

By John C. Draper

The principal object of an underground coal mine transportation sys- tem is to move coal from the face where it is produced to the outside of the mine where it is prepared for market. In addition to the movement of coal, a transportation system must also move supply materials to the inside; it must move men in and out of the mine; and it must move rock and other debris out of the mine for disposal. Usually, mine haulage is divided into three categories named for the various areas of the mine served : 1) Pace Haulage. This is used at face from the mining machine to the intermediate transport system. 2) Intermediate or Secondary Haulage. Generally this is used between face haulage and main haulage. Sometimes this is merely an extension of the main haulage. 3) Main Haulage. This extends from the secondary haulage to the outside and is usually either a track or conveyor system, sometimes a combination of both. Transportation modes of coal in the various areas pose different problems, but the whole system must be synchronized. For instance, at the face the coal should be removed quickly from the various places to facilitate the movement of machines, whereas the intermediate system just passes the same coal on to the main transportation system. Selection of the transportation system must be based on the seam height, mining methods used to extract the coal, capacity required, ease of coal handling, and haulage efficiency. When driving rooms, mine cars can be used. They often simplify supply haulage and rock handling because they can also handle these items. Conveyors could be used and would have a greater haulage capacity but they cannot handle supplies easily or segregate the rock. When using a different mining method, such as longwall, mine cars cannot be used because there is not sufficient space to move the cars at the face nor can the cars haul the coal without delays which would cause excessive losses in production. As for ease of handling at a longwall face, conveyors are convenient because of their low height, ruggedness, and ability to operate while being advanced. Similarly, in low seam heights, conveyor systems are easier and more economic to operate than the regular track haulage because roof has to be taken down for handling mine cars of adequate size.

Jan 1, 1981

By Earl L. Rau

At the present time, whether an electric utility uses high-sulfur coal is largely determined by geographical location, clean air legislation, and the age of the power plant. Acid rain legislation may well be the new force to change the use of high-sulfur coal. A number of coal papers were presented at the SME Fall Meeting in St. Louis, MO, Sept. 1986. These papers discussed high-sulfur coal in relation to its economics, trends and prospects, value of cleaned coal, and coal-cleaning technology. The Clean Air Act of 1970 required the reduction of SO, emissions from most stationary electric-generating plant to 0.5 kg (1.2 lbs) SO2, per million Btu of heat input. It was possible for many plants to utilize lower-sulfur coal or a mixture of high-sulfur and low-sulfur coal to meet this emission requirement. This 1970 act greatly increased the use of low-sulfur western coal. The 1978 amendment to the Clean Air Act required all new plants to use Flue Gas Desulfurization (FGD) to meet a requirement of 70% reduction of potential SO2 emissions to 0.3 kg (0.6 lb) SO2 per million Btu of heat input. With the requirement to use FGD, there was little or no demand for low-sulfur western coal in new eastern power plants. If' new acid rain legislation is enacted, it likely will require sizeable reductions in SO2, emissions by existing coal users. There might then be a shift to greater use of low-sulfur coals. The low-sulfur coals could come from either western or eastern mines or from cleaned eastern coals. Economics will dictate the approach older power plants will take to meet new regulationsuse FGD or low-sulfur coal. Coal use for electric power generation has changed from 1970 to 1985. Table 1, (Land, preprint 86-333) shows the change. The table shows that in 1985, US utilities used about 255 Mt (280 million st) of high-sulfur coal. Some 45 to 54 Mt (50 to 60 million st) of high-sulfur coal is being used in plants equipped with FGD and most of this use is in eastern states. This leaves 200 to 209 Mt/a (220 to 230 million st) of high-sulfur coal being used without FGD treatment. These plants are at risk if acid rain legislation is passed. Most of these plants are in the east and are more difficult to supply with low-sulfur coal. Estimates range from $20 to $30 billion capital investment and from $5 to $9 billion dollars per year operating costs to reduce SO2, emissions from those old medium-to-high-sulfur coal-burning plants. G. W. Land, (preprint 86-333), also presents tables relating to power generating costs and coal types. The author's tables are extracted from Land's tables IV and V. Table 2 here presents costs for a new plant. This includes the impact of different heat rates, fuel costs, fuel qualities, operating and maintenance costs, labor costs, and waste disposal costs. Table 3 compares estimated costs to retrofit an older plant with a limestone scrubber versus switching the same plant to low-sulfur eastern or western coal. Costs of switching include upgrading the electrostatic precipitator and - with western coal - costs of derating. Bhagwat, (preprint 86-325), details the trends in high-sulfur coal markets from 1974 to 1984. He also gives some prospects for high-sulfur coal markets, 1984 to 1994. The prospect depends mostly on whether acid rain legislation is enacted. By 1994, if New Source Performance Standards (NSPS) continue to be applied, high-sulfur coal demand from six mid-west and eastern states is expected to increase by 27 Mt (30 million st). • L. A. Khan and J. W. Baxter, (preprint 86-373), discuss several coal cleaning technologies such as froth flotation of fines, bulk oil separation and oil agglomeration, oil agglomeration/flotation, and air-floc flotation. ? (Space limitations preclude listing references. A complete list is available from the author.)

Jan 1, 1987

By D. A. Stanley

Researchers at BCD Technologies and the University of Alabama have applied three statistical techniques to the information contained in a large database of measurements from a phosphate ore processing plant. The rust, Principal Components Analysis (PCA), replaces many measured variables with a few fundamental and un-correlated variables. The second technique, Projection to Latent Structures (PLS), is an extension of PCA that explores the relation between variables identified as either independent or dependent. The third technique, Cluster Analysis (CA), is applied to the results of the PCA or the PLS calculations, and identifies individual data clusters that represent uniform operating conditions in the plant. All three statistical methods can be seen as geometrical operations on the data viewed as points in space. A Viewer gives a corresponding graphical presentation of the data points. This paper describes the use of the above three techniques and the Viewer to explore part of the phosphate flotation database.

Jan 1, 1998

I found the article most interesting and timely. I consider it imperative that mine case studies like this one are undertaken and published so that the practicing mining engineer and geologist can obtain a better idea of what is the preferred estimation and grade control technique to use, given the site-specific geology, grade distribution, etc. Mining companies, with operating gold mines, need to be encouraged to either initiate such a study in-house or support a research project by financing a graduate student at a local university. The latter method is preferable since it not only provides research support but also an excellent training opportunity for young mining engineers and geologists. Ideally, what is needed is exploration drill-hole, blasthole, and mine production data that can then be used to assess the efficiency of different estimation and grade control methods by comparing the results obtained using the drill/blasthole data with the actual production data. It is hoped that more mining companies will realize that the rewards from this kind of research greatly outweigh the sometimes perceived problem with releasing data to a university. I noticed that the authors used the blasthole data to represent the actual values, the validity of which relies heavily on the two key assumptions with regards to blasthole assays and in-pit ore-waste selection procedures. Could the authors comment in more detail on these assumptions and what criteria they used to check the validity of the assumptions? Since the article was published, have the authors had the opportunity to compare their estimates with the actual production data? It would be interesting to compare other estimation methods with OK and IK. For example, how well did the polygonal method used at the mine compare with the results given in the paper? I would like to mention that B.L. Kwa completed a similar study (Kwa and Mousset-Jones, 1986) using exploration drill-hole, blasthole, and production data from the Alligator Ridge mine in Nevada. The study compared a variety of estimation methods, including IK, and the final results are shown in Fig. 1. The IK method exhibits the smallest difference between estimated and production results and is closely followed by the indicator moving window (Ind MW) method. Reply by Y.C. Kim We do concur with Professor Mousset-Jones with respect to the merit of such case studies performed at a local university. In answer to his request for comment on the validity of the two key assumptions made in our paper, we can only say that these are reasonable assumptions that are very difficult to validate. The assumption of blasthole assays being accurate can be partially validated by a careful analysis of the nugget component of blasthole variograms. The assumption that the in-pit ore-waste selection procedure in use was an effective one can only be validated through comparison between the mine estimates of ounces of gold versus the actual recovered ounces of gold at the mill. Neither one of these two assumptions has been validated to date due to several reasons. For example, the mine went through a change in management since the study. Similarly, no compelling reasons have been presented to perform detailed variogram analysis of the blasthole data. We do, however, believe that the obtained results (i.e., indicator kriging performs better than OK in a highly skewed deposit) will remain the same in relative terms, regardless of how valid these two assumptions may be. The most recent com¬parative study by D. S. Lindsey of the Montana College of Mineral Science and Technology also shows a similar ranking between ordinary kriging versus probability kriging, which is an extended version of indicator kriging.

Jan 1, 1987

By M. M. Botz, F. W. DeVries

Discussion by F.W. DeVries Comments on natural cyanide attenuation. The paper by Botz and Mudder in the November 2000 issue of Minerals and Metallurgical Processing had one disappointing aspect. Table 1 (p. 229) failed to include a significant species - neutral copper cyanide (CuCN). This compound has been most trouble-some in many plant-scale cyanide cleanup processes, includ- ing acidification-volatilization- reneutralization (AVR), hy- drogen peroxide cleanup and air/SO2 cleanup. It has also plagued many pilot efforts to use ion-exchange (IX) resin technology. This author elucidated the concept that, under some conditions, copper cyanide failed to precipitate out of treated effluents and appeared to exist as a filterable colloid that reported as copper and cyanide, with the latter measured as WAD (weak-acid-dissociable) cyanide in that analytical method. The concentrations measured were very near the levels of 0.2 ppm, often the regulatory limit for WAD cyanide. The copper concentration, using the 1 : 1 ratio for this com- pound, would then be more than 0.4 ppm, which could also be unacceptable. We discovered that it was unsafe to assume that the Cu(CN),- would surrender both cyanide ions simultaneously and allow cupric copper to be removed subsequently as, say, cupric ferrocyanide (Hatchett's Brown). This compound could be ignored as safely as the "iron blues," as essentially an inert solid, stable in geologic time if not exposed to sunlight. Under poorly controlled oxidation of the cyanocuprite ions (copper with four, three or two cyanides complexed to the copper), CuCN could well be formed and could not be assumed to precipitate out as a solid. The authors of the paper, by failing to include the CuCN species in their model, may have left users of the model prone to ignore the possible occurrence of the colloidal(?) form in the discharge, thus finding an effluent that failed to meet regulatory limits. (See addendum below.) This somewhat ephemeral material may well have been responsible for some large-scale failures of cyanide destruction technology to succeed at sites in North America and elsewhere. Two specific references alluding to this occurrence may be found: Griffiths et al., 1987, "The detoxification of gold-mill tailings with hydrogen peroxide," Journal of the South African Institute of Mining and Metallurgy, September, p. 282. DeVries, 1989, "Improved understanding in detoxification of cyanide leach liquors using peroxide" Rand01 Gold Conference, Sacramento CA, p. 15 1. It is advisable for workers in this area of cyanide removal or recycle to avoid assuming that CuCN will, with certainty, precipitate out of solution, or that the dicyanocuprite ion will part simultaneously with both cyanide ions and allow cupric copper to be dealt with by other reactions. One further piece of work may be incomplete. That is, a definitive study of the long-term behavior of CuCN that does precipitate when exposed to other compounds present in neutralized tailings. Addendum. The extremely complex reaction of cupric ion on cyanide- which, of course, results in oxidation of cyanide to cyanate, probably through the intermediate stage of forming cyanogen [(CN)2] from oxidation of HCN with its hydrolysis to cyanate and cyanide ion - involves two cyanide moieties and the reduction of the cupric ion back again to the cuprous state. This, then, can form the CuCN potential "solid" that was already the undesirable species. In cyanide cleanup, with either air/SO2 or hydrogen peroxide, copper is generally employed as the catalyst, so it is unclear what happens when the CU(CN)2-1 is ultimately exhausted. Certainly, the field evidence is quite clear that all

Jan 1, 2002

By C. G. Anderson

Extraction of valuable resources from complex orebodies will become more common. This is obvious for two reasons. First, the cleaner and more readily processed ores have continually been depleted through centuries of mining. Second, new technologies have evolved for effectively handling complex ores. One of these new processes has been in place at Sunshine since 1984. Based on sulfuric acid oxygen pressure leaching with nitric acid as an oxidizer, this patented technology allows silver and copper to be efficiently recovered from a complex sulfide concentrate. As efficient as it is, this process represents novel technology found nowhere else on an industrial scale. As such, a research program was undertaken in 1989 to optimize this leach process and to better understand it. From this study, several potential plant improvements were found and recommended. As a result, two of these options, nitric acid replacement by sodium nitrite and concentrate regrind are now used in day to day operation at Sunshine. Patenting of these changes was recently accomplished (1). Thus, the development, perceived mechanisms, plant integration and potential applicability of this technology will be covered in this paper.

Jan 1, 1994