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By T. M. Plouf, G. A. Lang

INTRODUCTION The finding or acquisition of ore bodies containing 200,000 opt gold and greater is becoming more difficult. There are more investors and small heap leach operators who have or can raise the capital to make the investments in the large ore bodies. Echo Bay Mines' approach has been to regionally develop small ore bodies, two or three in the same geographic location, which when considered together contain over 200,000 opt gold. CONCEPT The "world class" ore body syndrome, which many large gold producing companies maintain, can restrict company growth if the finding or acquisition of these large ore bodies is not forthcoming. Another approach has been taken by Echo Bay Mines. The concept is to regionally develop two or three small ore bodies suitable for processing with portable equipment. The systematic scheduling of mining, processing with a single flexible processing plant using the same management/ operator team can reduce your overall investment. The initial selection of mining and crushing equipment suitable for sun11 open pit operations is essential. A very viable option in the mining sequence is to use contract mining/crushing/screening/agglomerating and stacking. This option also lowers your initial capital investment and increases the flexibility in mining as you move from one mining operation to another. LOCATION AND SEQUENCING One single small ore body of itself is not attractive enough for the large producer. The recovery of the initial capital investment, although very rapid, does not make a viable project. The only alternative is to find two or more small ore bodies close enough geographically to each other so that the same management/operator team can be used to run the operations as each is put into operation. Echo Bay Mines has regionally located several ore bodies ranging in recoverable ounces up to 100,000 oz. These small separate ore bodies, located within a 30-mile radius of each other, for planning purposes have been considered as one single ore body. The first small ore body has the initial burden to recover the capital invested in the office, laboratory, power plant and processing plant. Other capital investments such as water supply, access and haul road, fencing, septic system, pads and ponds are reoccurring at each mine site. The first small ore body must recover the initial capital investment in the movable structures as well as the reoccurring capital investment in the fixed structures very rapidly. A total capital recovery time of 6 to 9 months should be used as an economic decision point for both choosing which small ore body should be processed first, and what the plant design capability should be. The sequencing of the small mines is very critical. All the fixed structures mentioned above must be installed at the second mine site before the first mine has been mined out. The ideal sequencing will allow for only one month between shutdown of the first small mine and production from the second. SPECIFIC OPERATION Echo Bay Mines has completed the construction of the Illipah Mine located in White Pine County, Nevada, with this small mine sequence approach. The Illipah Mine is a 40,000 ounce recoverable gold ore body. This ore body met all the necessary requirements. The Illipah project could support the capital investment in the movable structures which will

Jan 1, 1987

By William Bogan, Mark Vaivoda, Richard Daniele

Many factors must be considered when performing engineering and construction work for a major minerals project. Dravo's Afton copper smelter in British Columbia illustrates some of the E&C factors considered when setting up this new operation. In March 1978, Afton Mines Ltd., a Tech Corp. subsidiary, began blister copper production at the world's first integrated copper ore mining, milling, and smelting operation to use Top Blow Rotary Converter (TBRC) process smelting technology. This key TBRC smelting element is licensed by INCO Ltd. to Dravo Corp. The TBRC process uses a single vessel of modular design. It rotates for mixing and provides complete processing capability, from concentrate through blister copper. The Afton facility is about 16 km west of Kamloops, a town in south-central British Columbia. To make its operation as economic as possible, Afton officials considered several options for the copper smelter, the first in modern times in the province. Inability to extract precious metals-primarily gold-from the concentrates ruled out hydro-metallurgical processes. The final choice was between a conventional electric furnace/ Peirce-Smith converter combination and the TBRC process. The TBRC process was chosen for its versatility in processing the various Afton concentrates and grades and its 40% lower capital cost compared to the electric furnace converter system. The TBRC smelter -including concentrate slurry receiving, drying, storage, smelting, off-gas cleaning, and copper casting-had an installed cost of about $25 million. In December 1974, a pilot program was conducted by Teck, Dravo, and INCO at INCO's Port Colborne, Ontario, research facility. The pilot program demonstrated the TBRC's ability to smelt, convert, and clean copper slags. After the successful pilot effort, Dravo performed additional research to find an effective fixation system for high grade, intermittent SO2 gases. Once the go-ahead was given, design and procurement for the Afton copper smelter was provided by the Denver Division, Dravo Engineers and Constructors. Commonwealth Construction, Vancouver, constructed all Afton facilities. Afton's single furnace smelter is designed to produce 25-kt/a blister copper in 363-408 kg cake form. The product is sold to England for electrolytic refining. The TBRC process is only one of several innovations proposed by Dravo and chosen by Afton to insure the most modern smelter in North America and meet all British Columbia's environmental restrictions. Conventional smelting methods use a separate furnace for each process step. The TBRC process employs but one furnace. The Afton TBRC vessel measures 6.4 m in length and 4.3 m in diameter. The furnace is supported on a set of motor-driven rollers. These permit the vessel to rotate at variable speeds to 40 rpm. Rotation provides good mixing, high reaction rates, and effective heat transfer from the refractory sides to the molten bath resulting in a low thermal gradient. The furnace can tilt through a full 360° for charging, firing, and pouring. Smelting at Afton is accomplished in sequential fashion. Flotation concentrate and limestone flux are initially charged to the hot, empty vessel. Using the water-cooled, oxy-fuel lance, the initial charge is smelted. Once sufficient molten bath is obtained and slagging completed, the metallic concentrate is charged. Materials are charged through the water-cooled, variable position additive chute mounted on the TBRC

Jan 10, 1980

By D. C. Torre

Since the early 1960s, there have been efforts to develop and produce reliable continuous haulage systems capable of taking the full output of a continuous miner and transporting it, in one continuous flow, from the face to the tipple. The majority of the successful continuous haulage systems were placed in operation in the early 1970s. These were all of the bridge conveyor type. They usually consisted of bridges and mobile bridge carriers (MBC) attached at the inby end to the continuous miner and at the outby end to the panel conveyor itself. These successful bridge conveyor continuous haulage systems are either of the belted or chain type. The most successful belt type continuous haulage systems are those manufactured and marketed by the Long-Airdox Co. Until early 1977, Jeffrey Mining Machinery Division manufactured and marketed the most successful chain conveyor type haulage sys¬tems. In 1977, Long-Airdox reentered the market with a chain-type system to complement its belted system. The original bridge-type continuous haulage system was introduced to the coal mining industry in the early 1960s by the old Long Co., a predecessor to the current Long-Airdox Co. These original units were of the chain conveyor type. In the early 1970s, there were also developments in continuous haulage, employing different material handling media. These in¬cluded the Serpentix, a flexible type belt, introduced by the Joy Manufacturing Co. There was also a serious attempt to introduce hydraulic-type continuous haulage by Consolidation Coal Co. and Conoco Research. Despite some startling successes, particularly with the bridge-type haulage systems, the general and widespread application of continuous haulage has eluded the mining industry. Some of the problems are related to underground conditions; others to people, both management and labor. One of the greatest obstacles to universal success continues to be the difficulty of accepting the philosophy as well as the machinery of continuous haulage. Nevertheless, efforts continue to spread the success of commercial and viable systems in the industry. There is also effort to improve on these systems, and on methods for their successful use. In addition, efforts continue, with private and public support, to develop new types of continuous haulage equipment and systems. Bridge Type System Developments Because they have already been successful in some US and overseas mines, new developments in bridge-type systems will first be touched on. The Long-Airdox Co. has continued to improve both its belted and chain continuous haulage system. In the belted system, the company recently introduced a new design that will be mainly applicable in higher coals and in noncoal mining applications, as shown in the accompanying figure. This new system is equipped with 305-mm-diam (12-in.-diam) pulleys in the bridges and mobile bridge carriers, in place of the 150-mm-diam (6 in.-diam) pulleys in the standard system. The modification is designed primarily to improve belt life and keep belt problems to a minimum. An operator, the Clinchfield Coal Co., a Division of Pittston Coal, developed the detaching of the bridge-type continuous haulage system from the miner by placing a mobile bridge carrier at the head of the string. However, Long-Airdox and Jeffrey have now made this an option to their at¬tached-type system. In the US, there are many chain type haulage systems detached from the miner with an extra MBC at the head. There are two such belted systems

Jan 7, 1982

By John W. Carson, David S. Dick

The last few years have produced new methods of characterizing flow properties of bulk solids. Abrasive wear and particle attrition can now be accurately measured in a laboratory and scaled up to field condi¬tions. In addition, the effect of large par¬ticles on a material's flow properties (cohe¬sive strength, bulk density, wall friction) can be determined as well as the effects of var¬ious additives such as freeze conditioning agents. New developments have occurred with devices to improve the reliability of solids flow. Much attention has been paid to find¬ing ways, by using inserts, to provide a mass flow pattern with relatively shallow hopper walls. Such devices can also provide reli¬able in-bin gravity blending of bulk solids or completely nonsegregated discharge. Other developments have occurred in the area of belt-to-belt transfer chutes and sealing screws. The technology and hardware for running bench scale flow tests on bulk solids samples and transforming the results into reliable designs has been in existence for nearly 30 years (Jenike, 1964; Marinelli and Carson, 1986). It has been shown repeatedly that reliable materials handling systems can be designed to avoid the problems of no-flow due to arching or rat-holing, erratic flow, segregation, loss of capacity, or spontaneous combustion. Running flow properties tests and carefully designing a bulk handling system are not expensive, particularly when compared to downtime costs and continual flow problems. New developments in characterizing flow properties of bulk solids include: • a shear tester for measuring the properties of run-of-mine ores containing large par¬ticles; • attrition testers to simulate a wide range of flow conditions; • a tester to measure the wear rate of liners in bins, feeder parts, etc.; • a dustiness tester to measure the potential for dust and the effect of dust control meas¬ures; and • techniques for measuring the strength of frozen ores and the effect of freeze condi¬tioning agents. In the area of improving the flow of bulk solids, there are new techniques for ensuring reliable flow without using up excessive headroom. These include: • belt-to-belt transfer chutes to handle po¬tentially sticky, abrasive, and dusty ores; • designs for the interface between a bin/ hopper and apron or belt feeders to avoid hangups and minimize feeder loads; methods to feed solids reliably into a pres¬surized process; and designs for handling systems that will mini¬ mize attrition, avoid segregation, or auto¬matically blend materials together. Characterizing flow properties Large shear cell It has long been known that the cohesive strength of a bulk solid is due mainly to its fine fraction. However, the effect of the larger particles (that were assumed to re¬duce the strength) could not be measured. The development of larger scale (205 mm¬diam and 0.6 x 0.9 m or 8 in.-diam and 2 x 3 ft) shear testers has made it possible to measure the properties of a wider range of materials. Now cohesive strength measurements can be made with up to 100 mm (4 in.) particles. The results obtained from the large shear testers, together with the results of the nor¬mal shear tests on the fine fraction, are used in the design of reliable stockpile reclaim systems and crusher feed systems. A high percentage of large particles can reduce the measured shear strength of a bulk solid. However, segregation may alter the size distribution and should therefore be considered (Carson, Royal, and Goodwill, 1986). In addition, larger particles will

Jan 1, 1989

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

It is imperative that the hydrogeologic frame- work and the ground water flow systems of a Potential mine area be investigated prior to initiation of mining. Such investigations are necessary so that potential mine water inflow problems can be anticipated and controlled. Obvious difficulties can arise if large inflows of ground water occur in underground workings. Costs of water removal and disposal, if not foreseen and included in economic planning, can severely limit or eliminate profit from a mining operation. Safety problems associated with large inflows are an additional potential problem. The inflow of water into a mine necessarily alters the associated ground water system. The inflow also may impact nearby surface water regimes, depending upon the interaction of the surface water and ground water systems. The inflow of water into a mine inevitably causes a decline in the ground water levels in the vicinity of the mine. This decline may be restricted to a small area or it may extend a considerable distance, depending upon the hydrogeologic properties of the associated hydrostratigraphic units. A decrease in surface flow will result if the area of ground water level decline intersects a stream or spring that is hydraulically connected to the ground water system. Mine inflow problems historically have been evaluated after-the-fact. Premining hydrologic investigations only recently have been initiated. Brown's (1977)* report on anticipated hydrologic impacts of oil shale mining in the Piceance Basin of Colorado is an example of premining hydrologic planning. Observed difficulties inherent in designing a data collection network and then utilizing the data to predict mine inflow have led us to the preparation of this manuscript. This manuscript reflects a two-phase effort. The first phase describes and develops analytical and numerical techniques for mine inflow prediction. Emphasis in this section is placed on the selection and application of data collection techniques and data analysis procedures. The second phase applies the prediction technique(s) developed during phase 1 to two existing mines. The objective of phase 1 efforts is to evaluate the hydrologic variables associated with surface water systems and ground water systems that can affect the inflow of water into an underground mine. Subsequently these variables are incorporated into predictive techniques for estimating underground mine water inflow and its effect on associated ground water and surface water systems. The specific objectives of phases 1 and 2 are to: 1) Determine the hydrologic parameters required for predicting underground mine water in- flow and the potential effects of the inflow on surface and ground water systems. The hydrologic parameters represent both premining and past initiation conditions. 2) Outline methods for obtaining the necessary input data for the predictive techniques in pre- mining and post initiation periods. 3) Present existing numerical techniques for predicting underground mine water inflow including assumptions Ad required input. 4) Evaluate existing numerical techniques previously tabulated based on their applicability to mine water inflow prediction. 5) Attempt to develop analytical techniques for predicting underground mine water inflow. 6) Develop, through modification of existing models, a general mathematical model as a numerical technique for predicting underground mine water inflow and the resultant effects on surface and ground water systems. 7) Apply the predictive techniques to two existing mines selected for this purpose. 8) Compare the predicted inflow rates to measured inflow rates at both mines. 9) Compare predicted impact on district hydrology to observed impacts where possible. 10) Evaluate the predictive technique.

Jan 1, 1986

By Steve O. Green

Introduction The US coal mining industry is in the midst of a serious economic crisis. To overcome increased operating costs, operators must take advantage of technological advancements in mining systems and equipment. A relatively new concept available to underground operators is that of extending the cutting depth for continuous mining machines (continuous miners). This system is referred to as deep-cut mining, and may be described as mining cuts to depths greater than the standard 6 m (20 ft). Productivity is increased through the reduction in continuous miner tram time since the machine stays longer in each working face. To successfully implement a deep-cut mining system, three problems must be overcome: Exposing personnel to unsupported roof, methane removal, and controlling respirable dust. Remote control cutting The solution to the unsupported roof problem is remote control. Recent advances in remote operation, especially in radio remote controls, have resulted in what is generally accepted as a safe and reliable tool. By installing remote control on the continuous miner, the operator and helper can now be stationed in a supported area. This allows for improved visibility while cutting coal. With this in mind, the limiting factor now becomes the position of the haulage equipment operator. North River No. 1 uses Joy 21SC shuttle cars and Jeffrey 4110 diesel ramcars. These cars are pulled against the bumper of the continuous miner. Thirteen meters (43 ft) can be cut with the shuttle car and 14 m (45 ft) with the Ramcar before the operator of the car is inby permanent supports. By loading away from the bumper, about 1.5 m (5 ft) can be added, bringing total cut depths up to 14.5 and 15 m (48 and 50 ft), respectively. Methane removal The second problem involves methane removal, and the line curtain-to-face distance. In other words, how to maintain the line curtain to allow sufficient ventilation to remove methane and not expose workers to unsupported roof. The solution, in most cases, is to use a blowing-type ventilation system. Unlike the exhausting system that, to be effective, must be installed within 3 m (10 ft) of the working face, the blowing system can be effective at greater distances from the face. Numerous studies have documented the advantages of a blowing-type ventilation system over an exhausting system. Furthermore, a blowing system can be significantly improved by in¬creasing the velocity of the air delivered without increasing the volume of air. With this greater ventilation efficiency, it is possible to secure ventilation plan approval from the Mine Safety and Health Administration (MSHA) for a greater curtain set-back distance. This eliminates the need for personnel to travel inby permanent supports to extend the face brattice. The distance from the face MSHA will approve, however, is on a mine-by-mine basis. MSHA will evaluate the particular mine's face methane liberation rates, along with the specific proposals regarding face ventilation volumes and velocities. Dust control The third problem is controlling respirable dust. This is actually a result of the solution to the second problem. In the blowing system, while mining on the slab cut, both the continuous miner and haulage units are located on the dusty side of the line curtain (Fig. 1). In the early 1970s, underground operators virtually abandoned this system, in order to comply with the 2 mg/m3 (0.87 gr per cu ft) federal dust standards. The blowing face curtain did offer increased face ventilation efficiency. With that, greater curtain-to-face distances were also

Jan 8, 1987

By J. D. Morgan

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

Jan 5, 1986

By Boris K. Krstev

INTRODUCTION The flotation process, in the real industrial complexes, is characterized by the existence of the significant number of the close reciprocal interrelationship between the input and output Darameters. Traditional complexity is appeared with regard to absence of the suitable a priori information for the physico-chemical characteristics of the flotation process. It’s led to needful of taking of consideration about the undefinition which have got the place in the real conditions and application of the statistical methods. The identification of the flotation process through statistical methods is conception of the more-stages procedure, including the collection and previous analysis, the choice of the model structure and valuation method, the estimation of the model un- known parameters and the results interpretation. The essential concept for the subject of this paper belongs to the empirical models ("black-box") and deals with optimizing in constant industrial plant process. The offered method envisages the application of process operation data in order to improve the work process or constant accomplishment of better results for the target function in relation to the former achievement. The optimum process solution isn't afixed value under the influence of various factors, but "a moving point" along the response surface and deviates from the scheme solutions. The evolutive planning method constantly "investigates and approaches" the optimum solution for the selected target function. As a matter of fact the expert operates the plant without interrupting the normal work process. The essential concept of method is to attain changes in the standard conditions, but so limited that they will not effect either production or the product. This method is based on statistical concepts and availability of following the risk of error. The possible influental factors for the flotation process (m=l, 2, 3) is following: the metal content in the feed ore; production level (t/h); grinding fineness in the flotation pulp; pH-values; the consumption of collectors and modifiers The possible target functions are: the recovery of plant capacity, metal recovery in the concentration process and concentration quality. The application of the evolutive operativity method in the chalcopyrite flotation- Bucim concentrator, combines the statistical data and the operator's experience in the interpretation and making a decision which adds up to the precise automatic performance of the process. The advantages of this method in relation to the existing procedure for the flotation process are the following: • The method is unexpensive and losses are reduced to the time necessary to collect the data and record them in the computer. This method provides a constant process control and a quality decision which is better than those brought by an ex- pert with great experience ("hit-and-miss method"). • The method also provides incidental data about the effects. • The method should not be considered as a procedure for solving an incidental problem ("crash program"). • The obtained results must be interpreted by an appropriate expert in order to be considered instructive. MATHEMATICAL STATISTICS Taking into consideration the complexity of the enrichment processes it's necessary the application of "black-box" and mathematical modelling. The presence of the indefinite elements compel us towards the both theory of the probability and mathematical statistics. The mathematical statistics by the mineral processing investigation of the valuable raw materials in the last years has had greater signifance and widely spread application. The application of the mathematical statistics methods for the technological pro- cess analysis and the construction of the mathematical models is bound up by the task putting of the automatic control system de- sign (Fisher R. A. 1941, 1949). The statistical analysis in the enrichment processes is applied formerly, especially in the sampling estimation. Contemporary, in the last years is imposed the tendency of mathematical statistics application for the obtaining of quantity characteristics and quantity valuation from the individual influental factors. Above mentioned tendency in the significant level is based by the mathematical statistics development as a mathematical discipline close connected with the technical cybernetics general joining to the investigated object as a "black-box". Another essential factor in the mathematical statistics development is appeared the created mathematical theory of experiments as a new, to some extent, autonomous statistical domain. The theory of experiments has demonstrated exclusively importance for the investigation of the complex more-factorial processes in the chemistry, mineral processing, metallurgy etc. The investigation by the enrichment processes of the valuable raw materials and their management is possible by the statistical method application. In the same time, it's possible to use the wide-spread traditional methods of the statistical analysis, such as the methods of the dispersion, correlation, regression etc. Also, it may be mentioned above all the methods of the statistical planning of the extreme experiments. The more perspective are representatived some cybernetic methods which are based by the probability as an approach in the process analysis, for example, the identify of the ways and forms, the graph application as an additional instrument in the correlation analysis, the dynamic programming etc. The mathematical statistics application in the laboratory investigation scale is connected above all by the experimented material analysis and the compact representation of the obtained results. The mathematical statistics methods are more varrying and may be applied in the investigations of the wide circle of problems

Jan 1, 1996

By George S. Austin

Perlite is rhyolitic glass with 2 to 5 percent by weight (wt.%) bound water held within the silicic-glass structure. It is often distinguished by vitreous, pearly luster and concentric perlitic fractures ("onionskin"). Occurrences worldwide are associated with Tertiary to middle Quaternary volcanic rocks. Perlite is in silicic domes, plugs, laccoliths, flows, vitric tephra, chill margins of dikes and sills, and welded ash-flow tuffs. It is a hydrated volcanic glass formed through secondary alteration of obsidian by the incorporation of water into the glass/silica structure. Post-emplacement hydration (weathering) of cooled glasses by meteoric water is the principal alteration process, but some hydration by magmatic or meteoric water during or shortly after emplacement, or hydration by water exsolved during cooling may occur. Commercial "perlite" is any volcanic glass that expands or "pops" when heated quickly to plasticity while evolving steam and thus forming a lightweight, frothy material. The expansion temperature typically ranges from 1,100 to 1,500°F. An expanded volume increase of 10 to 40 times the original volume yields expanded densities of 1.5 to 11 pounds per cubic foot (lbs/ft3). Perlite is mined by eight companies at 10 operations in six western states. New Mexico accounts for more than 80% of the total tonnage mined. The rest comes from Arizona, California, Colorado, Idaho, and Nevada. In 1992, domestic production was 571,000 short tons (st), domestic apparent consumption was 591,000 st, and the value of crude perlite was $15.3 million. In 1991, with about the same production and price of crude perlite, the gross value for expanded perlite was $101.7 million. Perlite exports in 1992, primarily to Canada, are estimated an 30,000 st; and imports, from Greece, are an estimated 50,000 st. In 1992, 70 plants in 33 states expanded crude perlite. Construction uses of expanded perlite accounted for 66% of total domestic sales. Expanded perlite used as filter aid accounted for 15%. All other markets, including use as a filler and in agriculture, accounted for 19%. Leading states for expanding perlite, ranked by order of sales, were Mississippi, Pennsylvania, California, Illinois, and Georgia. Domestic perlite encounters transportation disadvantages in some areas of the eastern United States compared to perlite imported from Greece. However, western U.S. perlite exports to Canada offset this in large part, and net import reliance for perlite is about 3% of domestic consumption. Perlite grab samples were tested at the New Mexico Bureau of Mines and Minerals Resources (NMBMMR) for their basic suitability in horticulture, construction, filter aids, and fillers/extenders using Perlite Institute techniques modified by the NMBMMR. Most commercial perlite ranged from 2.5 to 3.0 wt.% bound water for all major types (classical, granular, and pumiceous). Expanded density varied widely, but the lowest density of 1.9 lbs/ft3 was for granular. Average brightness of expanded perlite was 47.6 to 69.3%. Nonexpansibles, commonly obsidian or "dead" perlite, ranged from 18 wt.% to less than 1 wt.%, with granular perlte generally the lowest. Compacted density was not related to either perlite type or nonexpansibles present. Sieve analyses of the -50+100 mesh product showed that, in general, pumiceous perlite produces more fines and classical and granular perlite average 60 to 80% -30+100 mesh after expansion. Samples were tested for crystalline silica that ranged from below the detection limit of 0.5 wt.% to nearly 0.2 wt.%, but most tested below 0.5 wt.%. Although quartz content can be lowered in the expanded product by the expansion process, producers can further lower crystalline silica in the product by mixing low-quartz crude with high-quartz perlite before expansion. Chemical analysis by x-ray fluorescence confirms the rhyolitic composition of perlite, but indicates that surface weathering may increase the concentration of some elements, such as potassium and calcium, by infilling cracks or fissures developed during or after perlite emplacement.

Jan 1, 1995

By Robert W. Lawson

INTRODUCTION The two basic types of mine-air compressors are the positive-displacement compressors and the dynamic compressors. Positive-displacement compressors con¬fine successive volumes of air within a closed space and increase the pressure by reducing the volume of that space. Reciprocating and rotary compressors operate on this principle. Dynamic compressors utilize an impeller to acceler¬ate the air, with the increased air velocity converted to pressure through a stationary diffuser. Centrifugal compressors and turbocompressors operate on this principle. Reciprocating Compressors Reciprocating compressors include both single-stage and multistage units. For mine service the most com¬mon is the two-stage intercooled compressor having an atmospheric pressure intake and a discharge pressure of 689 to 861 kPa (100 to 125 psi). The cylinders of a reciprocating compressor may be either single acting or double acting. In the single-acting cylinder, compression occurs on only one side of the piston, while, in double¬acting cylinders, compression occurs on both sides of the piston. Fig. 1 shows a cross-sectional view of a typical two-stage double-acting water-cooled reciprocat¬ing compressor. Fig. 2 illustrates the sequence of opera¬tions occurring in the cylinder of a reciprocating com¬pressor. Most modern mine compressors are double acting and are driven by electric motors. Induction motors are used on low-powered compressors, and synchronous motors are used on larger compressors. Steam-driven compressors once were common and still are applied where a reliable source of low-cost steam is available. For this type of compressor, steam is a desirable power source since a governor that senses the discharge pres¬sure can be used to control the steam inlet. This can be used to regulate the compressor speed to compress only air sufficient to meet the mine's requirements (up to the capacity of the unit). Unlike a constant-speed electric¬driven compressor, no other regulation is required for the cylinders of a steam-driven unit. Rotary Compressors Rotary compressors include sliding-vane, liquid¬piston, two-impeller straight-lobe or cycloidal, and helical-lobe units. Each of these has specific advantages and applications. Of these, the helical-lobe or "screw" compressor has the greatest application for supplying "100-lb air" [689 kPa (100 psi) ] to small- and medium¬size underground mines. Fig. 3 illustrates the operation of a sliding-vane compressor. Fig. 4 illustrates the operation of a helical lobe or screw compressor. Fig. 5 illustrates the operation of a two-impeller straight-lobe compressor. Fig. 6 illustrates the operation of a liquid¬piston compressor. Centrifugal Compressors Centrifugal compressors deliver 0.94 m3/s (2000 cfm) or more of air, and they are best suited to supply¬ing large volumes of compressed air for base loads. The efficiency of a centrifugal compressor declines rapidly when its operation deviates from its design point. If the power cost is not a critical consideration, a centrif¬ugal compressor is attractive as a result of the low initial and installation costs and the high degree of reliability. Fig. 7 illustrates a typical four-stage centrif¬ugal compressor, and Fig. 8 is a cross section of this type of unit. COMPRESSOR SELECTION The process of selecting a compressor requires sev¬eral considerations, including the capacity, the control system, the location, the power supply, and the cooling provisions. Compressor Capacity To determine the required capacity, the total air quantity and pressure requirements must be calculated. The air consumption of various tools, as listed in Table 1, must be included in this calculation. Some of the other considerations include: 1) The correction factor for altitude, as listed in Table 2, must be included in the calculation. While pneumatic tools require additional compressed air as the altitude increases, the compressor produces less. 2) A correction factor must be incorporated for the use of multiple tools. Table 3 lists appropriate correc¬tion factors for rock drills.

Jan 1, 1982

By Gary Beerkirkcher

INTRODUCTION The mine hoisting equipment that is selected and installed at a mine is for the lifetime of the mine and, therefore, it is extremely important that a proper choice be made. Several types of mine hoist designs, including double drum hoists, single drum hoists, friction hoists, and various special designs of these, are available to¬day. Each has advantages and disadvantages and they should only be used in suitable applications. Drive systems for mine hoists must be evaluated together with the mechanical systems for each hoist application. It does not make sense to properly select the hoist mechanical system and then choose a drive system that turns out to be a bottleneck. MINE HOIST TYPES There are two basic types of mine hoists available in the world today: the drum hoist, on which the hoist rope is actually stored during the hoisting cycle, and the friction (Koepe) hoist, which merely passes the rope over the wheel during the hoisting process. Drum Hoists Single Drum Hoist: A single drum hoist (shown diagrammatically in Fig. 1) can be used as a service or production hoist with the cage or skip in balance with a counterweight (Figs. 2 and 3). It can efficiently service one or more levels, since the location of the counterweight at any time is not important. As a pro¬duction hoist with skips in balance, the single drum hoist is best used for single level hoisting. All rope adjustments for proper spotting must be done manually for the single drum hoist. This has to be done periodi¬cally to compensate for rope stretch. A variation of the single drum hoist is the divided drum hoist. If multilayer winding is necessary, the single drum hoist must have a divider to allow a sep¬arate compartment for each rope. If a counterweight is used with a divided drum application, the counter¬ weight rope can be wound on a smaller diameter (Fig. 4). Consequently, it moves a lesser distance than the main conveyance and rope adjustment problems are re¬duced. Double Drum Hoist: A double drum hoist in oper¬ation is shown in Fig. 5. Double Drum Hoist With One Drum Clutched: As a service hoist with cage and counterweight, a double drum hoist with one drum clutched (Fig. 6) can serve several levels efficiently. The advantage of having the clutched drum is that it allows rapid adjustment of ropes due to initial stretch. Some users may prefer the single drum hoist with manual rope adjustment, while others consider the added expense of the second drum and clutch is justified to make rope adjustments quickly. The value of this feature for a particular application should be discussed with the hoist manufacturer. As a production hoist with skips in balance for a multi¬level operation, the clutch can be adjusted for efficient hoisting from any level. Double Drum Hoists with Both Drums Clutched: There are also double drum hoists that have both drums clutched (Fig. 7). The main advantage claimed for this type of hoist is that if something happens in one of the two compartments, the hoist can operate in the other compartment to raise and lower men and sup¬plies. This feature is particularly favorable if there is only one shaft entrance to the mine. Multiple Drum Hoist: The multiple drum hoist, a South African innovation, is a type of drum hoist about which everyone should be aware. Basically, it uses two hoist ropes per conveyance which at great depths per¬mits use of smaller ropes and smaller drums. There are none in use in the United States today. However, they should be considered in very deep mines. Friction (Koepe) Hoists This hoist (Fig. 8) was introduced to the mining world by Frederick Koepe in 1877. The basic prin¬ciple of the friction hoist is that the hoist rope passing

Jan 1, 1982

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Variations of cut-and-fill stoping are the most com¬monly used methods for mining narrow, nearly vertical veins in the United States. These methods offer the best results for maximum recovery with minimum dilution of ore from veins that are irregular along either the strike or dip. The vein is developed along the strike length in 61 to 122-m (200 to 400-ft) blocks or stopes and vertical lifts of 48 to 61 m (160 to 200 ft). The level development is either on the vein or parallel to the vein with crosscuts intersecting the vein at right angles. Chutes, manways, and haulageways are established in the center of the stopes at the lower level of the ore block. Mining then proceeds by blasting down horizon¬tal cuts of the vein in both directions from the chutes. As the blasting continues to the outer limits of the stope, the broken ore is slushed into chutes. When a complete horizontal cut has been made, the service raise, including chutes, is raised to a height of a few feet from the back and the stope is filled with either sand fill or waste rock. Then the mining cycle is re¬peated until the stope reaches the level above. LEVEL DEVELOPMENT PRACTICES FOROVERHAND CUT-AND-FILL STOPING The two basic methods used are: the development of the haulage drift within the vein, or development of the haulage drift paralleling the vein in the foot or hang¬ing wall with crosscuts intersecting the vein perpendicu¬larly at 61 to 122-m (200 to 400-ft) intervals. The il¬lustrated methods of development show some of the variations which can be used to develop a stope block. Development for cut-and-fill stoping does not require the more elaborate schemes outlined for shrinkage or sublevel long-hole stoping. Many factors must be considered before deciding on a particular development plan and usually the nature of the vein or host rock will dictate which specific method to use. If the vein is relatively straight, then develop¬ment on the vein is feasible. Excessive ground pressures will necessitate the leaving of a sill pillar (Fig. la) and highly fractured or "heavy" ground will require timber¬ing (Fig. 1b). Development on the vein has the advan¬tage of reducing capital outlay and waste handling prob¬lems since the mined rock has value. But removal of the sill pillar is costly and time-consuming. Also, where no sill pillar is left, excessive timbering and subsequent level repair is a drawback. Development in either the footwall or hanging wall is the more common procedure and offers greater flexi¬bility in overall development of consecutive stoping blocks by partially disassociating the level development from stoping (Fig. 2). Chutes, manways, and service compartments com¬prise the next stage of development and are established in a variety of ways. They provide access to the stope, ventilation ducts, and muck removal. Several different development systems are in common use and average costs for each have been developed. Raise borers are used to establish ventilation and ac¬cess from level to level. They are normally reamed with 1.5 or 1.8-m (5 or 6-ft) cutterhead and may either be left bald or lined with a 1.2-m (4-ft) diam corrugated liner plate, depending upon ground conditions. The use of an Alimak raise climber is another effec¬tive method of driving a raise from level to level. The size usually is 2 m (7 ft) square. This will accommodate square-set timber which is later installed. This method of raising is not common, although it is a valuable tech¬nique especially when level spacing is several hundred feet.

Jan 1, 1982

By J. C. Agarwal

Major scientific discoveries cannot be programmed. New or frontier technology, however, can and should be planned. It is the vital force underlying the business strategy of a vigorous and profitable mining and metals company. Generally, a radical departure from existing technology is required to 1) Enhance the competitive value of ore reserves, i.e., to turn low grade or marginal reserves into competitive reserves. 2) To increase yield and decrease costs. 3) To decrease investment. 4) To increase mine life. 5) To improve product quality to meet market demands. The very definition of frontier technology implies that it is so different that the old plant, if already in use, will not be directly suitable to use the new technology. Therefore, a priori, we know that the commercialization of frontier technology would very likely involve new investment. The recommended technique of economic and technical analyses, I propose to call "Success Analysis.'' Simply stated, it forces us, by step- by-step procedure, to ask "what is needed to succeed in this particular project." Therefore, the first step is to identify what is meant by "success." The definition of "success" requires an agreement with top management or sponsors of the project as to what constitutes "success" in their minds as contrast to the technical people perception of "success." In general, top management would define success as meeting: 1) The technical objectives, i.e., technology being developed would produce the desired product at the acceptable quality. 2) Financial objectives, i.e., when commercialized, the project's capital cost, operating costs and return on investment are better than the projects they would normally fund. The return has to be better because new technology is inherently riskier. 3) Marketing objectives in that the project will produce product of acceptable quality in timely manner, i.e., scale-up to commercial plant would not cause undue delays in start-up and that large fix-up expenditure would not be required to protect the projected return on the investment. Many projects are started in the laboratory to develop "Frontier Technology" in mineral processing. Most of these efforts are never commercialized. There are many causes of these failures. The major causes are listed below: 1) Inadequate communication with the sponsors of the project so that project objectives are defined and agreed to by all. 2) Ill-defined technical problems. 3) Incomplete development of technology. This has two major components: a. Incorrect scale-up and engineering. b. Inadequate start-up and maintenance planning. 4) Incomplete and over-optimistic process economics. 5) Lack of understanding of the future behavior of the marketplace. 6) Failure to complete R&D effort by the time it is needed. 7) Incorrect projections of human resources required; that is, technical, managerial and operating. 8) Lack of realistic forecast of the expenditures required for commercialization of the R&D effort. These pitfalls can be avoided by comprehensive planning and the use of success analysis technique. The major elements in "Success Analysis" are: 1) Identification of the "need" for new technology. 2) Unbiased and complete scale-up engineering. 3) Acceptable economic feasibility with enough margin of error in the sale prices and operating costs to justify the inherent risks in frontier technology. 4) Timing for availability of investment capital and markets. Many failures in the development of frontier technology are preordained because the initial selection of the research project has been without adequate attention to the needs of the industry. Many very innovative hydrometallurgical processes for recovery of copper from concentrates were developed in the past twenty years. They have not adapted except to a very minor extent because they did not satisfy the success analysis criteria. The copper industry in North America needed a process to compete against high grade ores and to avoid

Jan 1, 1985

By J. A. Marshall

"INTRODUCTION The purpose of this paper is to investigate the potential for use of UAVs in underground mines and present a prototype design for a novel autorotating UAV platform for underground 3D data collection. The mining industry has recently shown increased interest in the use of unmanned aerial vehicles (UAVs) to assist in everyday operations [20, 11, 4]. Above ground, small UAVs are in some cases a more efficient, inexpensive, and safer alternative to manned aircrafts currently used for photography, inspection and security [6, 10]. For example, attaching camera, infrared and LiDAR payloads, UAVs can provide a low-cost method of obtaining highly accurate 3D photogrammetric data and aerial photography. UAVs are now commonly used in open pit mining operations for applications that include stockpile surveying, 3D pit modelling, facilities management, accident reporting, progress monitoring, and environmental assessment [11]. There are numerous companies offering UAV services for mining applications; e.g., UAV Geomatics, Leica Geosystems, Sky Futures, SenseFly and FlyTerra, to name just a few. Common services include stockpile and open pit volume computations, environmental assessment, aerial mapping and photogrammetry. Although the aforementioned services are thosemost commonly provided by UAV surveying companies, they may also add value to other daily mining operations. For instance, SenseFly offers pre and post-blast monitoring in order to identify the presence of misfires and wall damage. This data can also be used to reconcile the blast results with expected results. GEM Systems offers a UAV equipped with a magnetometer for mineral exploration surveys. Companies such as Barrick have even used UAVs for solids modelling at tailings dams and stability monitoring [11]. UAVs in mining have so far been mostly limited to surface applications. Harsh underground environments pose many obstacles for flying UAVs. The confined space, dampness, reduced visibility, air movement, and lack of control signal propagation hinders most operators from being able to fly a drone underground. It may be that truly practical uses for UAVs underground will require either autonomous or semi-autonomous flight capabilities. Although there are many difficulties with flying underground, the potential benefits from a working system could greatly improve mining operations. The potential benefits of deploying UAV platforms underground include access to unreachable and dangerous locations and removing personnel from unsafe situations. These benefits have the potential to greatly improve mine monitoring and mine safety. Research has shown that current UAV technologies exist that allow for autonomous indoor flight. Extensive research has been done to develop UAV systems that are capable of performing on-board simultaneous localization and mapping (SLAM), which can allow them to navigate and map a foreign environment autonomously [5, 7, 1, 13]. Grzonka [7] successfully used an open hardware quadrotor to autonomously navigate and map an office building. The research outlines the localization, mapping, path planning, height estimation and control of the autonomous quadrotor. Other research has been done that exploits autonomous UAVs for search and rescue. Kassecker [9] proposed a software and hardware framework for a quadrotor capable of indoor and outdoor urban search and rescue and Rudol [19] developed a system for human body detection and geolocalization using an autonomous UAV.The use of autonomous UAVs in search and rescue has the potential to improve situational awareness and surveillance for a rescue team. Similar research has been done that uses autonomous UAVs for indoor exploration. Pravitra [15] outlined a strategy for autonomous exploration of miniature air vehicles (MAVs) within indoor environments and Rudol [19] developed a method for using an unmanned ground vehicle (UGV) and a UAV for cooperative indoor exploration. These studies show that aut"

Jan 1, 2017

By August V. Castelli

Drilling fluids have changed over the years from water and/or water frequently mixed with local clays to scientifically com- pounded materials (mud). This change has occurred due to changes in drilling methods from percussion to rotary: depth maximum from 21 m in the first oil well to more than 9 000 m; and variety in types of well or hole to gas, oil, water, and mineral exploration. Environmental concerns about the composition of the fluid as to heavy metals and oil on the cuttings has recently caused changes in the materials used in the mud. The drilling fluid should perform the following functions (Darley et al., 1980): 1) Carry cuttings from beneath the bit, transport them up the annulus, and permit their separation at the surface. 2) Cool and clean the bit. 3) Reduce friction between the drilling string and the sides of the hole. 4) Maintain the stability of uncased sections of the borehole. 5) Prevent the inflow of fluids-oil, gas, or water-from permeable rocks penetrated. 6) Form a thin, low-permeability filter cake which seals pores and other openings in formations penetrated by the bit. In addition to these, the fluid should: 1) Not injure drilling personnel nor be damaging or offensive to the environment. 2) Not require unusual or expensive methods of completion of the drilled hole. 3) Not interfere with the normal productivity of the fluid- bearing formation. 4) Not corrode or cause excessive wear of drilling equipment. A number of industrial minerals, chemicals, and other materials are used as additives in the drilling fluid to accomplish these functions. Industrial minerals make up the highest percentage of the additives in a drilling fluid. Drilling fluids are classified according to their principal components: water, oil, and gas. If the major component is either water or oil, the component along with the suspension of solids, cuttings and additives is referred to in the trade as a mud. The largest use of some industrial minerals, such as barite and bentonite, is in drilling muds. For other materials, such as gypsum and limestone, drilling usage is only a small percentage of total demand. RAW MATERIALS [Table 1] lists the industrial minerals used in drilling fluids, their form, specifications, and uses. Each raw material will be discussed as to its mineralogy, physical and chemical properties, and the location of major deposits with respect to drilling fluids. Attapulgite Attapulgite is a crystalline hydrated magnesium silicate clay mineral, prized for its viscosity in saline waters. The particles are needle-like rather than the platelets common to most clay minerals. The specific gravity is 2.5 and the Mohs hardness is 1 to 1.5. Attapulgite is one of the materials sometimes referred to as fuller's earth. The major producing areas in the United States are southern Georgia and northwestern Florida. It also is mined in Spain, Senegal, the Republic of South Africa, Turkey, India, and Australia. Barite Barite is crystalline barium sulfate. It has a high specific gravity for a non-metallic mineral, 4.5 when pure, and is used as a weighting agent. Generally, the specific gravity ranges from 4.20 to 4.35. It is relatively soft with a Mohs hardness of 3 to 3.5 and is not very abrasive. Barite is inert, non-toxic, and clean to handle. Common contaminants are quartz, iron minerals, gypsum, and calcium carbonate. If there is considerable gypsum in the barite, it will thicken the mud. The major deposits in the United States are in north central Nevada, near Battle Mountain. The major producing countries, other than the United States, are China, India, Mexico, the former USSR, Turkey, and Morocco. The major deposits in India and Nevada, and some in China, are stratiform, commonly quite pure bodies in bedded sedimentary rocks.

Jan 1, 1994

By Richard R. Klimpel

INTRODUCTION Froth flotation is the most widely used and economic means of concentrating metal sulfide ores such as those containing copper, lead, zinc, nickel, molybdenum, and pyrite. Also recoverable are other metal species that are often associated with sulfide ores, such as cobalt, platinum, gold, silver, etc. Flotation of metal sulfide ores has been successfully practiced at the industrial scale for more than 50 years. It is the relative technical/economic ease of floating metal sulfide minerals that has been a major driving force behind the rapid expansion in the supply/demand of metals over this 50 year period. It is estimated that 1.6 billion tons of feed sulfide ores are processed each year in the free world based on the froth flotation process, Klimpel (1987) and the manufacture/use of associated chemical reagents is a major portion of the 1.5 billion dollars (U.S.) free world mining chemicals business, Klimpel (1988b). The use of froth flotation appears to be increasing partially due to the ever decreasing feed grades and liberation sizes of the feed minerals. Froth flotation has shown itself to be a flexible process that lends itself well to many solid/solid separations. Thus, with surprisingly little equipment modification, separations can involve very different particle sizes and densities, and relative weight ratios of materials to be separated. In addition, the flotation process is economic, especially when compared to size reduction, its associated precursor process. Flotation also lends itself to continuous operations with a variety of equipment configurations. Such operations exhibit an ability to vary feed rate of solid to the process by as much as 50% without a total collapse of separating efficiency. In addition, the widely used mechanical flotation cell is scalable from 2.8 to 57 or 85 m3 (100 to 2000 or 3000 cu ft) with surprising ease relative to other unit engineering operations over the same relative size increase. The separating medium used is water. Most of the chemicals required -- pH regulators, frothers, collectors, activators, and depressants -- are all relatively inexpensive and common. They are not usually used in large quantities. While there has been a tendency for the quality of metal sulfide ores in any one geographic location to deteriorate over time due to mining intensity, the overall global supply of sulfide minerals is still very ample. Thus, the physical supply side of sulfide ore concentrates has not been under strong technology improvement pressure with increasing metal demands. Part of this is due to the ability of capital to consistently move to new geographic sources having suitable-grade sulfide ores. In these moves, the actual froth flotation technology used (including collector chemicals) has had to change little. As an illustration of this trend, of the approximately 80,000 metric tons of thio collectors used commercially (1980) in the free world, almost 987 were known and manufactured in some form 25 years ago. This is clearly not typical of reagent development in other process-related industries. In addition, the industrial-scale practice of froth flotation applied to sulfide ores has proceeded since the 1920's with often little direct (predictive) scientific explanation due to the extreme complexity of the flotation process. Empirical testing has been a mainstay of industrial flotation reagent development and use. Even today there is often strong disagreement between researchers as to the mechanisms of chemical flotation practices that have been performed successfully on an industrial scale for many years. Thus the industrial process of froth flotation, especially as applied to the recovery of sulfide minerals, poses a dilemma: such a process is widely used, economic, versatile, forgiving, etc. and yet is still not very well understood in a mechanistic sense leading to prediction of results from fundamentals even after 50 years of usage. Lack of such predictive ability has not limited the general over-all industrial use of froth flotation in sulfide mineral recovery. Rather the lack of

Jan 1, 1989

By B. P. Boisen

INTRODUCTION Application Measurement of rock deformation is essential to stability and behavioral monitoring of underground openings. The difficulties of obtaining adequate sample distribution, density, and orientation can frequently be overcome by the use of single and multiple position borehole extensometers. In its simplest form, a borehole extensometer consists of an instrument head, usually anchored in the collar of a borehole, and a number of in-hole anchors, each of which is fixed in place at a known distance from the collar. As rock adjacent to the borehole deforms, the component parallel to the borehole direction causes a distance change between adjacent in-hole anchors, as well as a distance change between individual anchors and the instrument head. These changes are measured by means of a mechanical link (usually wire or rod) extending to the instrument head from each anchor. The resulting data can be used to compute the distribution, magnitude, rate, and acceleration of deformation in the material intersected by the extensometer's borehole. Wire Vs. Rod Extensometers Since the early 1960s myriad varieties of borehole extensometers have become available through the combination of various special features of two basic instruments, the wire and rod versions of borehole extensometers. Recent publications have recommended shallow depth limitations for wire extensometers, and favor rod extensometers for a broad range of borehole inclinations and depths. There are three main reasons for using a deep wire system: its cost effectiveness; the fact that tremendous depth means tremendous rod weight which will cause the rod to buckle or rub against the borehole or casing wall; and users' attitude and proper understanding of how to install and apply a wire-type extensometer. Here experience plays a big role. Over the years, many have learned the hard way that untensioned rod extensometers are simple and inherently inexpensive, but limited to vertical up holes where sensitivity is not influenced by friction between the rod elements and the borehole wall or protective conduit, the inertia of the rod elements, or the direction of the movement to be measured. In deep, horizontal, or inclined holes, and in instances where compressive as well as tensile strains or displacements are anticipated, tensioned rod or wire systems are preferred. System Components Borehole extensometers are made up of at least the following five principal components: the instrument head; the collar anchor assembly; the downhole system of anchors and rods or wires; the signal cable if the head is equipped for electronic readout; and the readout apparatus, whether mechanical or electronic. These components are shown in Fig. 1. Borehole Requirements Most extensometers are designed for installation in nominal 75 to 57-mm (3.0 to 2.25-in.) diam drill holes. Either a diamond drill or a percussion drill hole is satisfactory, provided the variation in drill-hole diameter is within the tolerance of the standard anchor design. The basic system may be modified for other hole diameters by changing the dimensions of the collar anchor and downhole anchor assemblies. Some instrument heads and collar anchor assemblies require enlarged borehole collars, though usually not over 150 mm (6 in.) for a depth of 1 m (3 ft). It is desirable, although not essential, that holes be free-draining. Holes must be clean and free of debris before installation. Depth/Orientation of Extensometer Boreholes Extensometers are designed to measure axial strain components in the instrumented drill hole. As the drill hole is axially deformed, the distances between individual downhole anchors and the instrument head change, as do the distances between adjacent anchors. The distance changes measured by a multiple position borehole extensometer, for example, are reduced to functions of time, hole depth, anchor position, or other parameters of significance in terms of the structure under investigation.

Jan 1, 1982

By John C. Franklin

INTRODUCTION Alpha-emitting radon daughter products are a recognized health hazard contributing to the development of lung cancer in persons exposed to excessive concentrations over an extended period. The primary control technique used by the mining industry is dilution with fresh air. Because the uranium is deposited in porous sandstone in most U.S. mines, the control of this hazard with dilution requires excessive volumes of air. As the mines become larger and deeper, the cost for ventilation becomes a major expenditure for the mining industry Most mines are maintaining levels of radon daughters lower than required by law but to maintain these present levels the mining companies are going to be drilling more vent holes or using other control techniques as the mine expands. These control techniques are being investigated by the U.S. Bureau of Mines to determine their effectiveness, cost, and safety for underground use. [Instrumentation] The Bureau's early research used batch sampling of both radon and radon daughters for determining the concentrations. In both cases it was determined that batch sampling did not give a complete understanding of what was happening to the concentrations. Therefore, continuous monitoring systems were developed for both. A continuous flow-through radon monitor was developed by the Bureau (Franklin, et al.., 1976) using a modified Lucas flask mounted on a photo-multiplier-tube (PMT). The air was filtered twice prior to entering the flask to remove dust, diesel smoke, moisture, and radon daughters. The signal from the PMT was then inverted, shaped, and driven to a data acquisition system (DAS) which accumulated the signals. The DAS has two comparators for setting the scan time and interval, a 24-hour clock with Julian date, a printer, and a perforated paper tape punch. Sampling time is determined by the type of experiment being conducted. Our normal sampling time is one 10-minute scan every 15 minutes, but could be set as low as one scan every minute or as long as a 16-minute scan every 167 minutes. There are two models of the DAS. One is a 10-channel system and the other a 30-channel system. On a 10channel system normally 6 channels are used for radon and radon daughters and the other four channels are used for temperature, absolute pressure, air velocity, and relative humidity as described by McVey (1977). Figure 1 shows a scan of five channels monitoring the radon concentration from intake to exhaust in a stope for a non-work day. If the radon concentration stayed this constant during work days, then continuous monitoring would not be necessary. [ ] Droullard (1977) developed a continuous alpha-beta monitor. This monitor has been modified so that it can be used with the DAS previously mentioned. Control Techniques The U.S. Bureau of Mines has conducted research in controlling airborne radon/radon daughter for 10 years. Major efforts have been conducted in sealing the rock to prevent the release of radon, bulkheading to seal nonproductive areas, backfilling with classified mill tailings, and positive pressure ventilation. Also, tests have been conducted to determine the mining cycles contribution to the radon problem. Surveys One stope with two intakes and one exhaust was continuously monitored to determine the effects of blasting, mucking, and drilling on the airborne radon concentrations (Franklin, et al., 1976). Two other detectors were located inside the stope which were close to the mining activity. In this stope no appreciable change in radon was observed during drilling holes for blasting and roofbolts. This study showed increases up to 40 percent in radon for slushing in the exhaust drift. The radon remained high and very eratic during slushing, as can

Jan 1, 1981

By M. C. Stuttle

MDL Rock Lasers has undertaken successful mapping trials of abandoned mine workings using its underground cavity scanning system. The work was being performed for Kalgoorlie Consolidated Gold Mines (KCGM), Australia's largest gold producer. MDL Rock Lasers supplies underground surveying systems to the mining and quarrying industry KCGM's Fimiston Super Pit, located in Kalgoorlie, Western Australia, is Australia's largest open-pit gold operation. Fimiston's final pit design is expected to be 3.7 km (2.3 miles) long, 1.5 km (0.9-miles ) wide and 540 m (1,770 ft) deep. During the early days of Australian mining, underground operations were labor intensive. Mining was selective and generally narrow stopes were mostly backfilled. Improved processing technology and mine mechanization during the 1970s and 1980s has permitted larger scale operations to adopt less selective underground-stope operations. The "super pit" concept was a logical step for mining companies working toward economies of scale. However, the industry recognized that there was no precedent for excavating a 500-m- (1,640-ft-) deep open pit through ground previously affected by old workings. To appreciate the geotechnical challenge facing KCGM, it is necessary to consider that under the pro¬posed area at the open pit, stoping reached a level of 1 km (0.6 miles) and was supported by more than 2,000 km (1,240 miles) of development headings. The company was fortunate because plans were available for the Golden Mile underground operations. The existing plans were interpreted by the mine's geotechnical team and computer models were constructed of the underground workings. These models can be imported into three-dimensional mine planning packages, such as the Vulcan's Unix-based software, Envisage, to assist analysis and design. This prior knowledge of the workings provided the focus for investigating and confirming ground conditions using probe drilling. After several iterations of drilling, followed by detailed analysis, several decisions were made relating to the nature of the ground and the mining approach to be taken. It was recognized that the stope models were not perfect due to hu¬man error, lack of original survey information and progressive deterioration in ground conditions. It also became apparent that the sole reliance on probe drilling was inefficient in terms of time and quality of information. This was particularly true in complex areas where stopes are in close proximity. A major concern was the presence of open stopes (voids). In many cases, stopes consist of combinations of filled and void sections. The condition of pillars within these underground workings is extremely important. In time, pillars collapse and voids will propagate in upward and lateral directions. So systems needed to be developed that allowed rapid and accurate verification of ground conditions. KCGM was introduced to MDL Rock Lasers while investigating technology for underground mining. The two companies decided to use the C-ALS, MDL Rock Laser's underground laser, cavity scanning system, to improve the management of mining through affected ground. This system was to be used in conjunction with probe drilling and void mapping activities. A field test in one area of the Fimiston Mine verified the capability of the C-ALS system to quickly and efficiently assist in this decision making process. The C-ALS scanner features lightweight, carbon fi¬ber alignment rods from Measurement Devices Ltd.'s (MDL) Boretrak MKII. This is a borehole deviation system that stops the two axis measurement head rotating and lowers the system down a borehole. The scanner is deployed by a 110-mm (4.3-in.) borehole, up, down or sideways. The operator can then carry out vertical and horizontal plan section scans at any user specified ARC (an angular increment of 2°) or the distance between two points on the ARC (CHORD) increments. Windows '95 software enables the operator to control the cavity scanner remotely, surveying the area in real time. The system makes data quick and convenient.

Jan 1, 1999

By James R. Arnold

Gold Fields Mining Corporation started processing ore from it's Chimney Creek operation in northern Nevada in November, 1987. Since that time production has averaged about 200,000 ounces per year from the milling and dump leaching circuits combined. The events that led up to commissioning of the plant started in January of 1985 when the Gold Fields exploration team drilled their first hole in an area now known as Discovery Hill. The major orebody was located later in the year in what is currently the major focus of activity, the North Pit. Subsequent drilling since plant startup has found another 14 million tons of ore two miles south of the original orebody. This South Pit orebody is currently being stripped and metallurgical testing is underway. Engineering Engineering on the Chimney Creek plant started in earnest in May, 1986. Davy McKee Corporation from San Ramon, California, did both the engineering and the construction management. Before engineering began the entire senior staff of the Plant Department was hired and assigned to the project. Senior staff included the Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist. Gold Fields does not employ an internal central engineering staff to design their plants. We feel that the people who design the plant should be the ones who will be required to run it once it has been commissioned. This philosophy affords a continuous accountability from start of design through commissioning and operations. All four of the senior staff spent the most critical phase of the engineering stage at Davy McKee, so the plant is truly a reflection of all four disciplines. In bringing on the Maintenance General Foreman from the start, there was a focus placed on ease of maintaining the plant. Another philosophy that was adopted from the start was that the plant be easy to work in and maintain for both operators and maintenance personnel. Numerous design criteria were given to Davy McKee to ensure this. For example, we required that every piece of equipment could be reached easily with a crane or fork lift. All conveyors were required to have walkways on both sides for ease of maintenance and a minimum of 30 inches of clearance under them for ease of cleanup. Visibility for operators was considered very important and the control rooms and general plant layout were designed with this in mind. A list of about 100 general rules were given to Davy McKee during the early layout meetings and these were held as the general design criteria throughout engineering and construction. While engineering was progressing in San Ramon, California, dump leach test work was being done at the site. Three 6 foot diameter fiberglass columns were constructed and filled to a 20 foot height with low grade ore. Leaching characteristics were determined and the dump leach design was based on information collected from these columns. Construction The major contractors on the project were The Industrial Company for the bulk of the piping and structural, EMKO on the electrical/ instrumentation and Hunewill for most of the earth work and leach pad construction. All plant operating and maintenance personnel were employed eight weeks before startup for intensive training. Instructors were Gold Fields personnel, including foremen, and Davy McKee engineers. On occasion, selected suppliers representatives were brought in to discuss major pieces of equipment. Project completion was targeted for January, 1988 at a projected cost of 78.9 million dollars. Actual startup was about 1 1/2 months ahead of schedule and slightly under budget. Full design capacity was achieved the first full month of operation. Crushing Run of mine ore is delivered to the crusher and dumped into a 90 ton hopper which is topped with a 30" X 30" opening stationary grizzly. Ore is fed to a 50" X 60" Telsmith jaw crusher with a 5' wide apron feeder. The closed size setting of the

Jan 1, 1989