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By Fred N. Kissell, George H. Jr. Schnakenberg

INTRODUCTION A variety of instruments are available for measuring or monitoring the performance of underground mine-ventilation systems. In general terms, the instruments may be classified as those that measure air velocity and those that measure gaseous concentrations. All costs herein are in terms of 1978 US dollars. The mention of a specific manufacturer or device is not intended to be an endorsement by the US Bureau of Mines. AIR-VELOCITY INSTRUMENTS The basic instruments used for measuring the air velocity in mines are the vane anemometer and the smoke tube. Vane Anemometer Of the air-velocity instruments, the 102-mm (4.0¬in.) vane anemometer is the most common and is available as either a low- or high-speed type. The low-speed anemometer generally is the most suitable for measuring the velocities in ordinary airways. For a rough check of the velocity in an airway, it usually is satisfactory to hold the anemometer by hand, positioning it in the center of the airway for 30 sec. However, the resultant error may be as high as 25% , and such a hand-held approach is unsuitable when accurate or reliable measurements are required. To obtain more accurate measurements, the proper procedure is as follows: 1) Since holding the anemometer by hand generally causes the instrument to read about 15% high, it is mounted on a 0.6-m (2-ft) extension rod. 2) The airway is divided into equal right and left halves. A 1-min traverse is used in each half, moving the anemometer smoothly up and down in a zigzag pattern so that the entire half is covered within the allotted minute. 3) The manufacturer's correction table is applied to the readings to adjust the velocity calculation as necessary. Whenever possible, anemometer readings should be obtained in a long straight section of airway that has a constant cross-sectional area. Bends and obstructions should be avoided, since they cause turbulence and other discontinuities in the airflow and can degrade the accuracy of the velocity measurements. Although a series of velocity measurements at one location usually corresponds to within a few percent, this is not an indication that the airflows calculated from those readings are completely accurate. One reason is that the correction table provided with the instrument generally is not for that specific instrument; instead, it represents the average correction for all such instru¬ments made by the particular manufacturer. Most cor¬rection tables specify a correction factor of from 0 to 15%, depending upon the velocity. However, even after correction, the instrument error still may range from 3 to 5%. At low velocities such as those below 0.76 m/s (150 fpm), the instrument error can be two or three times greater than this, ranging from 6 to 15%. The new ball-bearing anemometers generally perform somewhat better at low velocities than did the older conventional anemometers. Another source of error is introduced when measur¬ing the cross-sectional area of the airway or entry. Under the best of circumstances, measurement errors, instrument errors, and a host of other minor errors all combine to cause a total error of at least 10% in the velocity calculation. The vane anemometer also can be used with reason¬able accuracy to measure airflows in mine-ventilation ducts. In this application, the anemometer is mounted on a rod and is held at the center of the duct end. For a duct that is discharging air, the average velocity in the duct is 85% of the centerline reading (Northover, 1957). For a duct that is taking in air, the average velocity is 70% of the centerline reading (Haney and Hlavsa, 1976). To measure the airflow discharged from a regulator or from a small hole in a stopping or bulk¬head, a correction factor for the area is necessary. A good approach in this situation is to traverse the area of the regulator or hole, holding the anemometer with an extension rod. This provides an average velocity that is multiplied by 85% of the measured area of the regulator or hole. In all cases, the manufacturer's instrument cor¬rection table must be used and applied properly. For accurate results, the anemometer should be returned to the manufacturer for periodic cleaning and checking. If it is in daily use, the anemometer should be returned about once per year, and proportionally less frequently if the usage is less frequent than on a daily basis. Smoke Tube The smoke tube may not appeal to individuals who believe that good measurement results can be obtained only with expensive, complicated, and fragile instru¬mentation. Nevertheless, smoke works about as well as anything for the routine measurement of low air velocities in mines. The following procedure yields reasonably good results: 1) Two marks are scratched 7.6 m (25 ft) apart on the floor of the airway. 2) The smoke tube is used to release a cloud of smoke in the center of the airway, about 0.9 m (3 ft) upstream of the first mark on the floor. 3) A timed interval begins when the leading edge of the smoke cloud passes over the first mark, and the interval stops when the leading edge of the cloud passes over the second mark. 4) A factor of 20% is subtracted from the cal¬culated velocity to determine the true average velocity, providing a correction for the centerline and for the spreading effect at the front of the cloud. Velocities calculated with the preceding method generally are accurate to within 10 to 15%. In some instances, the cloud from a conventional smoke tube

Jan 1, 1982

By Roman Z. Pyrih

Subsurface migration of cyanide is attenuated by naturally occurring geochemical and biogeochemical processes. Geochemical mechanisms which remove or immobilize cyanide and metal-cyanide complexes include precipitation/coprecipitation, sorption, and volatilization. Geochemical properties of soils and geochemical materials that promote the functioning of these mechanisms include the following: 1) acidic soil pH; 2) presence of illite and smectite clays; and 3) presence of iron hydrous-oxides and organic carbon. Naturally occurring soil microorganisms play perhaps a more important role in attenuating cyanide movement than previously realized. Biodegradation of cyanide is a "renewable" process, limited in its effectiveness by the avalibility of oxygen and micronutrlents (usually phosphate) to the microorganisms.

Jan 1, 1990

By R. W. Barbaro, R. V. Ramani, K. V. K. Prasad, P. T. Luckie

Discussion by A. Unal Barbaro et al. (1990) implemented a tedious study aimed at delineating the short-scale spatial variability of sulfur in a coal seam. It is not possible, however, to extend their conclusions to any other coal seam or use their in any other fashion, because the background geology was not presented. Nor were the conclusions accompanied by geological interpretations in the paper. In addition, an unfortunate printing error occurred where the total sulfur (determined by High Temperature Combustion) variogram in Fig.3 was duplicated in Fig. 2 instead of the seam height variogram. A more serious error, however, has been committed in the definition of the variable, seam height. What is defined as seam height by Barbaro et al. (1990) is, in essence, mining height, and is not a regionalized variable that should be studied by geostatistical methods directly. Despite the fact that the variable, seam height, is not defined in the paper explicitly, the following quotation discloses the fallacy: "All roof rock that exceeded 1.8 m (6 ft) from the floor was not taken because the longwall was operated to not mine more than 1.8 in (6 ft) unless the coal seam height exceeded 1.8 m (6 ft)." From this definition, the seam height, and therefore the sampling height, is equal to the actual coal thickness, if the coal thickness is greater than 1.8 m (6 ft). Otherwise, it is equal to the sum of the coal thickness plus the thickness of the roof rock that complements the thickness to 1.8 m (6 ft). In the latter case, the seam height is constant and equal to 1.8 m (6 ft). It is possible to conduct a variogram study on a pool of samples that are realizations of two different variables. But the conclusions derived would not belong to any one of the two variables uniquely and, therefore, do not possess any significance. Geostatistical analysis is irrelevant for the sum of multiple regionalized variables formed by arbitrary selections. In a two-seam setting, for example, the mining height, as defined by the thickness of one seam at one location and the thickness of both seams at another location (due to quality and/or minimum thickness considerations perhaps), should not be used in the calculation of one common variogram. The two seams should be modeled separately. They can then be combined according to the specific purposes of the study. On the other hand, if a constant is added to a regionalized variable (to incorporate dilution perhaps), the variogram of the new variable will not change. Barbaro et al. (1990), surprisingly, does not give any geological interpretation of their results despite the fact that most of them can be explained by the origins of sulfur in a coal seam. The presentation of the results of a geostatistical study with no reference to the geology of the deposit is uninformative. It may also be misleading for the potential users of geostatistics, It is not unusual to find nuggets of pyrite in coal seams. In such cases, pyritic sulfur will probably display a spatial structure for only a very small distance that will appear in the experimental variogram as no spatial correlation. This very well known phenomenon is called the pure nugget effect in geostatistics (Journel and Huijbregts, 1978) and perhaps can explain the lack of correlation found for pyritic sulfur content. The lack of correlation found for the total sulfur content may also be explained in the same way because the total sulfur content is dominated by the pyritic sulfur content in this case study. One should notice, however, that the situation may completely be reversed after cleaning the coal. Not all of the inorganic sulfur should be expected to be in the form of pyrite nuggets in a coal seam. It may also be disseminated in the coal seam and it is expected that it follows a certain spatial structure. However, an existing spatial structure may be masked by including a part of the roof rock, rich in sulfides, into the seam thickness in an arbitrary fashion because areas having a sandstone roof sometimes are known to show a higher sulfur content due to the downward percolation of solutions rich in iron sulfides (Clark, 1979). Plants use sulfur in their growth processes. Much of this sulfur is bound organically during peat accumulation and coal formation (Cecil et al., 1978.) This suggests a spatial structure of some sort for the organic sulfur. However, it is not possible to test this hypothesis because the results obtained by the authors for the organic sulfur content are not given in this paper. For this reason, the conclusion that simple average of the nearby samples would provide the best unbiased estimates is questionable for organic sulfur and is not based on any substantial supporting evidence. It is suspected that no spatial structure was detected and this was due to high sampling and laboratory analysis errors. Before concluding that sulfur variability in the seam at the location of study was random, a more detailed study for the disseminated non-pyritic sulfur should have been conducted, not for the sake of scientific curiosity only, but also due to its utmost importance with regard to coal cleaning and emission control. Pyritic sulfur can be cleaned to a considerable extent, whereas organic sulfur can not, making the emission control strategies highly dependent on the spatial distribution of the organic sulfur (Knudsen, 1981). In the light of these facts, one wonders why Barbaro et al. (1990) did not present the results of their study for the sulfate and organic sulfur content. References Barbaro, R.W.. et al., 1990, "Short-Scale Spatial Variability of Sulfur in a Coal Seam,' Mining Engineering, Vol. 42, No. 11, pp. 1267-1269. Cecil, C.B., et al., 1978, "Geology of Ccontaminants in coal," report prepared for Environmental Protection Agency, North Carolina, 123 pp. Clark, W.J.. 1979, "An interfluve model of the upper Freeport coal Bed in part of western Pennsylvania," unpublished MS thesis, University of South Carolina, 57 pp. Journel, A.G., and [Huijbregls], Ch. J., 1978, Mining Geostatistics, Academic Press, London, 600 pp. Knudsen, H.P., 1981. "Development of a Conditional simulation model of a coal deposit," unpublished PhD dissertation, The University of Arizona, 109 pp.

Jan 1, 1992

By K. J. Henderson

INTRODUCTION Shrinkage mining at Inco Ltd.'s Crean Hill mine can best be defined as a horizontal breasting method utilizing the broken ore for the mining floor and to provide wall support, with draw from a series of boxholes or draw¬points at the base of the stoping block. In choosing to shrink a given block of ore, the initial investigation centers on the ore configuration and grade, the expected ground conditions in the stoping area, the expected profit compared to alternative mining methods, the future production requirements, development lead time, the availability and type of equipment in the area, and the availability of experienced miners. These factors are weighed against other mining methods and shrinkage mining is chosen when the advantages outweigh the dis¬advantages. These will be discussed in more detail later. At Inco Ltd.'s Crean Hill mine, shrinkage mining makes up 17% of total mine production and is decreasing due to the advent of advanced blasthole techniques, notably in-the-hole drilling. Shrinkage is an adaptable mining method that is currently employed in narrow irregular ore zones, as undercuts for in-the-hole blast¬hole stopes, in ore bodies that do not extend from level to level, and in areas where quick production is required with a minimum of capital outlay or minimum develop¬ment. Two major types are currently in use which are classified on the basis of type of draw system employed. These are the utilization of load-haul-dump (LHD) equipment drawing from drawpoints, and the slusher and train method drawing from boxholes into a slusher trench and loading into a train. A third method being phased out is drawing from chutes directly into a train. In all three methods the actual stoping operation remains the same. In the initial planning of a mining block, the ore body is first located using an exploration diamond drill¬ing program. This is essentially done at a 130-m (400¬ft) spacing and it gives an indication of the amount of ore, grade of ore, the configuration of the ore body, as well as other geological features such as ground condi¬tions which will affect future mining. This drilling is done from an exploration drift 165 m (500 ft) long in the hanging wall and covers a vertical extent of 330 to 670 m (1000 to 2000 ft). However, these data are not accurate enough for detailed planning. When the de¬cision has been made to mine, a preliminary develop¬ment drill drift is driven in such a position that it can serve the dual purposes of allowing the ore body to be drilled off in detail, or of later serving as a haulage or access drift. Diamond drilling from this drift is usually on a 15-m (50-ft) grid spacing and gives adequate information for detailed planning. Utilizing the drill data, a decision is made on the mining type. For a shrinkage operation, the stoping block is ideally 49 m (150 ft) in length and a maximum of 13 m (40 ft) in width. Lengths of 65 m (200 ft) are considered a practical maximum because of the length of time the back of the stope remains open during the min¬ing of a cut. A minimum length of 13 m (40 ft) is used because the amount of time spent in opening a new cut increases in comparison with the total stoping time. The maximum width of 13 m (40 ft) is used because ex¬perience with the ground indicates that generally stope backs tend to be unstable over greater widths. However, widths of 33 m (100 ft) have been mined. A minimum width of 5 m (15 ft) is used to allow an effective draw, but in unusual cases, a width of 3 m (10 ft) has been mined successfully. The dip of the ore should not be less than 1.04 rad (60°). Angles flatter than this will not pull properly resulting in inefficient mining because of ore hang-ups on the footwall side of the stope and the need for large stagings on the hanging wall. In considering shrinkage, mining ground conditions play an important role. Strong shearing, cross faulting, or dikes are avoided. These usually cause dangerous mining conditions and an alternative safer mining method which will give adequate support to these adverse ground conditions is employed. Production requirements are taken into considera¬tion when planning a shrinkage mining program. In it¬self, shrinkage mining is capable of producing ore in a relatively short time with little capital development. If quick production is required, this method can produce ore rapidly as all that is required is that the draw system be completed and the miners with jacklegs and stopers produce a limited amount of ore. The obvious disadvan¬tage is that not all the ore broken can be pulled from the stope immediately. However, a mining horizon with sev¬eral shrinkage stopes in different stages of production can adequately maintain scheduled tonnage require¬ments, the cycle being stopes under development, stopes actively mining, and stopes completed and pulling. When the decision has been made to use shrinkage mining in a given block, the haulage and draw system is driven. The proximity of the orepass system dictates whether the system is to be a LHD operation or a slusher trench operation. If the orepass system is in close prox¬imity to the stoping block, LHD equipment is employed because of its ability to operate efficiently over short haulage distances. A long tram to the orepass usually dictates a slusher trench operation with haulage by train. Internal orepass systems are used effectively with LHD equipment with a load-out and retramming taking place on a lower horizon. In addition, stoping blocks in older areas of the mine are developed track because of the capital cost involved in converting an existing level to trackless mining. DRIFT DEVELOPMENT USING LHD EQUIPMENT The haulage drift for a draw system using LHD equipment is driven parallel to the ore body in the foot¬wall at a sufficient distance to allow an 11.5-m (35-ft) drawpoint to be driven into the ore (Fig. 1). The actual position of the drift depends entirely upon the ore con¬figuration and the drift size depends on the equipment to be used. Drift sizes of 4.6 x 4.1 m (14 x 121/2 ft) for an St-4 and 3.8 x 3.6 m (111/2 x 11 ft) for an St-2 meet company and government clearance standards and allow sufficient space to accommodate a ventilation pipe. The

Jan 1, 1982

By G. Rainville, I. Servi, F. Katrak

Despite the severe drought conditions that reduced farm requirements for industrial mineral products, most industrial minerals markets in 1988 continued their growth or, at worst, remained flat. Earlier projections of output declines did not materialize in most segments. Preliminary estimates of demand in Europe and Asia show strong growth for most industrial minerals. Profitability in industrial minerals in North America was best in those minerals that had a significant export market or were not strongly regional. Growth and price trends in the more regional industrial minerals markets in the US such as sand and gravel, crushed stone, and cement, more closely followed the broadly disparate regional economic conditions. For example, sand and gravel and crushed stone grew strongly in the Pacific Coast and northeastern markets but not as much as in the Southeast. Total growth for sand and gravel in 1988 is projected at about 3% above 1987 levels. Combined production is up about 27 Mt (30 million st) to 2 Gt (2.2 billion st). Following the relentless trend of foreign acquisition of US cement companies (currently about two-thirds are owned by foreign interests), several aggregate operations have been purchased by foreign companies. Notable among these in 1988 was the acquisition of Rinker Materials Corp. by CSR of Australia. The acquisition of this Florida-based aggregate and concrete operation will expand the current holdings of CSR in the southeastern US, with operations consolidated under the Rinker logo. In addition, Pike Industries of New England and J.L. Shiely Co. of Minnesota were significant aggregate producers that were acquired by foreign firms. New England's only cement manufacturer, Dragon Products, was acquired by a subsidiary of Cementes del Norte of Spain. Dravo continued to expand its influence in the lime and limestone markets. It became the major supplier of construction aggregates on the inland river system with its purchase of Cyprus Minerals' limestone aggregate operations in Kentucky, Louisiana, and Texas. Although lime production continued to grow from 14 to 15 Mt (15.7 to 16.7 million st) in 1988, lime imports decreased for the fifth consecutive year to 145 kt (160,000 st). In the more export-oriented industrial minerals markets, performance was generally very good for 1988. Soda ash enjoyed an excellent year, with its price up to $102.50/t ($93 per st). This reflected the tight market situation for soda ash, particularly in late 1988. Soda ash production in 1988 was 8.6 Mt (9.5 million st), reflecting the industry's improved efficiency. Particularly significant was the increase in caustic soda prices that led to increased substitution by soda ash. The export market remained at 2.1 Mt (2.3 million st). Phosphate production recovered to the 42 Mt (46 million st) level, a 12% increase despite a soft export market. The price, however, remained soft throughout the year. W.R. Grace sold its interest in its Florida phosphate mine and its phosphoric acid complex as part of its divestiture of the agrichemicals business. The strengthening of the major producers has continued as lower cost capacity has been idled. Future permitting of phosphoric acid facilities and development of reserves will be necessary to maintain current production levels beyond the mid- to late 1990s. Despite new developments worldwide in the titanium minerals market, strong demand has continued to apply pressure to price, with concentrate and slag prices going up. The demand for high quality slag as feedstock for pigment production has resulted in process improvements in South Africa (Richard's Bay) and in plans by Canada to import high quality ilmenite by 1991 to produce a 90% TiO2 slag. Although growth in industrial silica sand applications was small in 1988, concentration in the industry continued. Unimin continued to acquire silica operations. Unimin is now the nation's leading producer of granular silica. The end users of silica have consolidated further. Owens-Illinois purchased Brockway Inc., a leading container glass producer. Three companies now control 75% of the container glass industry. ECC continued to be an aggressive purchaser of industrial minerals operations throughout the world. It acquired Cyprus Minerals' calcium carbonate business as well as two operations in Italy. In addition, ECC continued its aggressive acquisition of kaolin (Australia) and aggregate producers. 1988 was a good year for industrial minerals markets worldwide. More importantly, though, it was a year that showed continuing consolidations of reserve ownership in the industry around the world. Barite AN. Castelli, Baroid Drilling Fluids Inc. US mine production of barite decreased 9.4% during 1988. Consumption (sold or used by grinding plants) increased by 37.9%. Imports are estimated to have increased by 21.2%. World mine production decreased by 9.8%, according to the US Bureau of Mines. The value of domestically produced barite, fob mine, decreased 4.6%, according to the Bureau. The declared value cif US port of all imported ground barite during the first 10 months of 1988 increased from $37.16/t ($33.71 per st) in 1987 to $37.92/t ($34.40 per st), according to Bureau figures. Nevada continued to be the leading producer of barite with 72% of the total, followed by Georgia and Missouri. The Bureau of Mines estimates 69% of US mine production was used as a weighting agent in drilling fluids. The other 31% was used in barium chemicals, glass, or as a filler. Most of the production from Missouri, Georgia, and Tennessee was used in the non-oilfield sector. Of the total consumption used by grinding plants and chemical manufac-

Jan 1, 1989

By Roland F. Beers

At East Union, Knox County, Maine, near the shore of Crawford Pond, an outcrop of basic rock has been known for more than a century. The occurrence has been regarded as an unusual type of peridotite-pyroxenite, containing disseminated and massive pyrrhotite with nickel, copper and cobalt in recoverable amounts. From the outcrop, the sulfide body had been mapped more or less continuously for a distance of two miles through geophysical methods of several kinds. This paper describes the methods in sufficient detail to evaluate their significance. A core drilling program has con- firmed the location of the conductor and offers a positive means of correlating the geophysical measurements with the physical properties of the ore body. The history of the exploration offers a unique opportunity to evaluate the method of multiple geophysical methods.

Jan 1, 1962

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

6.1 INTRODUCTION A surface structure will suffer damage when the additional stresses induced by ground deformations associated with surface subsidence, plus the original stress introduced by construction de¬sign, exceed the strength of the structural elements. In this con¬text, there are two methods available for preventing and control¬ling surface structural damage: one is to strengthen the structure and the other is to design the mining operations such that ground deformations at the structure site can be reduced to an acceptable level. Mining operations include panel layout and mining tech¬niques. These methods are detailed in this chapter. It must be noted that most prevention/control methods men¬tioned in this chapter are used in the countries where the reference papers are cited. In the United States, the coal operators are not required to take those measures mentioned in this chapter. Some of the methods described in this chapter cannot be implemented with¬out changes in the current mining practice as permitted by laws. In addition cost of implementing those methods are not considered here. 6.2 PANEL LAYOUT As shown in Figs. 2.9, 2.10, and 2.11, permanent ground deformations in a subsidence basin mainly concentrate near the edges of the underground opening, and can be divided into four zones. A structure located in different zones will be subjected to different types and magnitudes of ground deformations. In laying out the panels, Table 5.1 and Figs. 2.9, 2.10, and 2.11 could be taken into consideration. Attempts could be made to avoid placing the structure on a location where the ground deformation to which that structure is sensitive is at its maximum. Therefore rational design of the panel is the simplest way to reduce structural defor¬mations. Panel design involves the determination of panel dimension, panel edge location, direction of face advance, and use of yield chain pillars. A. Panel Dimension Since longwalls in the US employ a multiple-entry system, where rows of chain pillars are left unmined, subsidence over those chain pillars is usually smaller. Therefore, whenever possi¬ble, the panel dimension could be designed such that a major structure or structures are over those unmined chain pillars, be¬tween adjacent panels, or some distance beyond both ends of the panels. At the center of the supercritical final subsidence basin, a structure will not be subjected to any final or permanent ground deformations. In order to create such a condition, the panel width must be such that the structure will be located beyond the major influence zone of the final subsidence basin, the minimum dimen¬sion of which must be: [ ] where L is the width or length of the final mined out gob, t is the width or length of the structure to be protected, h is mining depth and [ ] s is the angle of full subsidence. B. Panel Edge Location Wherever there is a permanent panel edge, there are large ground deformations induced on the surface on both sides of the permanent panel edge. Therefore whenever possible the panel di¬mension should be designed such that the permanent panel edges could be located in the areas with the least impacts. In terms of permanent edge location, it is best to eliminate any permanent panel edge under a structure or groups of structures. If this cannot be done, the panel should be lengthened to reduce the number of permanent panel edges, or narrower multiple panels advancing in the same direction in a staggered manner could be employed. If the structures are located in Zones II and III, the longer dimension of the structure must be parallel to the nearest perma¬nent edge (Fig. 6.1). But in Zone IV, the longer dimension should be tangential to the corner of the permanent panel edge. If the structure is inclined to the permanent panel edge, it will be sub¬jected to twisting and shearing. C. Direction of Face Advance The direction of face advance should be parallel to the long axis of the structure. But if the structure is to be located at or close to the center of the final subsidence basin, the direction of face advance should be perpendicular to the long axis of the structure. Careful choice of the direction of multiple face advance is the most effective way to reduce structural deformation and thus dam¬age. This applies the principle of overlapping and cancellations of ground deformations, due to multiple face advance, at the right time and at the right intensity, e.g., opposing tilts, concave and convex curvatures, tensile and compressive strains are induced simultaneously on the structure to be protected by two or more faces. D. Use of Yield Chain Pillars In US longwall panels there are generally two or three rows of stiff chain pillars between the panels. The combined width of the chain pillars ranges from 100 to 350 ft(30 to 107m). depending on mining depth. In general, surface movement above the chain pil¬lars after the panels on both sides have been extracted is much smaller, as compared to that in panel center. Thus in order to create critical or supercritical width of opening and eliminate sur¬face bumps over the chain pillars, yield chain pillars may be em¬ployed (Jarosz and Karmis, 1986). However if yield pillars are to be used, it must be designed such that it yields totally right after the panels on both sides have been extracted. Unfortunately cur¬rent yield pillar design techniques cannot predict when and how much it will yield. In summary, whenever possible attempts could be made to lay out the panel in such a way that surface structures are located above chain pillars between panels or above solid coal beyond both ends of the panels. In those areas the surface structures will most likely be unaffected, or if affected, the damage is so minor that no remedial measures are necessary. 6.3 CONTROLLED MINING TECHNIQUES Several mining techniques are available for reducing the sur¬face ground deformations of specific types. Regardless of tech-

Jan 1, 1992

A panel discussion on carbon-in-pulp processing of gold and silver ores was one of the highlights of the 1981 AIME Annual Meeting in Chicago. The session generated considerable interest and discussion among panelists and the audience. For those unable to attend the panel, the program was recorded and the discussion appears in this two-part report. The panel was co-chaired by Robert S. Shoemaker, vice president of San Francisco Mining Associates, and Laurence D. Hartzog, principal engineer with Bechtel Civil and Minerals Inc., San Francisco, CA. Panel members included: George Potter, consultant, Tucson, AZ; Kenneth B. Hall, metallurgical superintendent, Homestake Mining Co., Lead, SD; and Donald M. Duncan, resident manager, Pinson Mining Co., Winnemucca, NV. There have been several types of carbon-in-pulp adsorption vessels in use, such as the Dorr-type agitator, the simple propeller agitated tank, the Pachuca tank, and, lately, the draft tube-type of agitated tank. Which one of these is best and why? Potter: On the selection of the most appropriate adsorption vessel, there is a considerable difference of opinion that stems largely from the fact that every ore is different. The mesh of grind, the pulp solids, and the apparent viscosity of the pulp as caused by its clay content and chemical conditions all differ. The traditional Dorr agitator with the slow speed center sweep and either peripheral or center column air lifts has worked quite well on the finer grinds of all minus 65 and probably 70-80% minus 200 mesh. Pachuca tanks have also been successful and they may be capable of handling a coarser feed than the traditional Dorr tank. One uranium mill, for example, handles minus 28 mesh sandstone ore in a Pachuca. The most recent development and one that commands consideration is the draft tube agitator in which there is a turbine which closely fits inside a draft tube. Velocity in the tube is carefully calculated to avoid undue shear and thus abrasion of the carbon. The objective in all cases, of course, is to mix the granular carbon, which is typically 6 x 16 mesh, very gently but thoroughly in a slurry with minimum carbon abrasion. I do not know that there is one outstanding choice just yet and I have been unable so far to get information on C-I-P service for ore grinds as coarse as 35 mesh. Duncan: The major advantage of draft tube-type is the ease of startup. At the Pinson plant we installed draft tube-type agitators but it is too early to quote experience with them. We also haven't operated long enough to determine just what our carbon loss is. The advantage, of course, in the draft tube design is that it requires only about one-third the horsepower input of a conventional agitator. If it has no other advantages, it has that. Hall: The type of vessel most suitable for C-I-P adsorption depends on the type of ore treated and the prevailing operating conditions. In most cases, a deep tank with turbine-type mechanical agitator and low speed tip velocities would be satisfactory. Turbine-type impellers give a positive type of agitation which assures optimum ion contact and reduces short circuiting. Thorough aeration is possible with an air sparge properly located. A mechanical agitator can easily be started up after an outage, but it is usually necessary to drain and wash out a Dorr rake-type or Pachuca before restarting. The turbine-type impeller requires more power, but maintenance costs are negligible if rubber covered impellers are used. Carbon losses are minimal. In smaller plants, Pachuca type agitators provide adequate aeration and agitation. Bob Polak, Occidental Minerals (from the floor): Mr. Hall, you favor the mechanical-type agitator with the preface that the proper design is critical. Could you elaborate on this for us? Hall: The most important thing is low tip speed to prevent carbon attrition. I think the impeller should be sized so that you get a good sweeping action toward the bottom of the tank, across its bottom, up the sides, and back to the center. The advantage of the draft tube is that you get a more positive agitation and you probably get improved aeration too. Polak: Do you know of anyone using an upflow mechanical agitator rather than a downflow unit? Hall: I think that Bob Wilson at Custom Equipment has designed one where he has located the sparge directly beneath the impeller. The air bubbles are broken up by the impeller as the slurry passes through it. The slurry flows upward, outward, down the sides, and, again, back to the center of the tank. I don't know that any tanks of this design are in commercial use. Hans Von Michaelis, Randol International (from the floor): A question to any of the panelists on flat bottom versus conical bottom Pachucas. Hall: I would say the conical bottom would be most suitable in most cases. Some plants have experienced problems with flat bottom Pachucas in that they have a tendency to sand in on the sides of the tank so only the center of the tank is active. Duncan: With our draft tube tank we placed an inverted cone in the center of the tank bottom and blanked off the bottom comer around the circumference of the tank to facilitate movement of the pulp. Other than that, as far as conical bottom tanks are concerned, I'm generally opposed to them because of the cost and height differential. Larry Kramer, Kennecott Minerals Co. (from the floor): Mr. Duncan, you mentioned that the draft tube-type could be agitated with about one-third the horse-power applied to a conventional propellor agitated tank. I have trouble trying to pin down that kind of number. Could you give a bit of rationale as to why that mechanism has a lower horsepower requirement? Duncan: I think it has to do with the fact that below the propeller you have straightening vanes. The pulp, which is flowing vertically downward, is turned and flows upward again without any recirculation. You have an inherently less horse-power requirement in that type of tank. The one-third horsepower requirement is a number I obtained from Lightnin, which supplied us with details of the draft tube. There is probably much more recirculation with a conventional impeller type that is wasted motion. Potter: Last week I saw some agitators in South Africa with short draft tubes and actual turbine-type impellers. This plant was handling 200 kt of ore per month. The tanks were flat bottomed, and the agitation appeared to be quite satisfactory.

Jan 8, 1981

By K. Thomson

The Mule Canyon gold deposit, located 14 miles ESE of Battle Mountain, Nevada, was discovered in 1986 as the result of a GFMC grass-roots exploration program. Drilling to date has identified six discrete orebodies which contain a demonstrated gold reserve of 8.487 million tons at an average grade of 0.136 oz/ton, totaling 1,152,000 ounces. Mineralization is associated with argillic alteration and silicification, localized along NNW trending high-angle faults within Miocene basaltic andesites. Trace element geochemistry at Mule Canyon is typical of upper level hot spring precious metal systems, displaying elevated Ag, As, Sb, Hg, and Se.

Jan 1, 1993

By Q. Xu

The adsorption behavior of gold in ammoniacal solutions on to activated carbon has been studied. The gold loading capacity of the activated carbon used in the system and the recovery efficiency by the adsorption were examined. The parameters affecting the adsorption of gold on to activated carbon included temperature, agitation speed, concentration of 'cation and anion as well as that of ammonia. It was found that the maximum gold loading on activated carbon used was about 10 kg of gold per ton of carbon. The recovery of gold by the adsorption from ammoniacal solutions was dependent upon the initial gold concentration in the solution. A recovery of 98% of gold was obtained on to the activated carbon under experimental conditions after 120 hour adsorption. It was also found that the initial adsorption rate was relatively fast compared to the whole period of the adsorption. The apparent activation energy for the adsorption was measured to be 15.8 kJ/mol. It was believed that the kinetics of gold adsorption on activated carbon was controlled by pore diffusion.

Jan 1, 1995

By K. Norris

The Amax Gold Guanaco mine is located in the Atacama Desert, approximately 260 km southeast of Antofagasta, Chile. The 6,000 tpd leach operation produces approximately 90,000 oz of gold, and 360,000 oz of silver per year. Production began in 1987, and the operation has been managed by Amax Gold since 1992. The mine has been plagued in the past by high operating costs and low production. Several significant changes have been made to the operations which have resulted in lower costs and higher operating efficiencies. A review of these changes is presented herein.

Jan 1, 1997

By D. Kang

The adsorption behavior of silver in ammoniacal solutions on to activated carbon has been studied. The parameters affecting the adsorption of silver on to activated carbon included temperature, agitation speed, concentration of cations and anions as well as that of ammonia. It was found that the adsorption of silver from ammoniacal solutions was dependent upon the initial silver concentration in the solution and other factors. The recovery of silver from ammoniacal solutions was found to be not so effective as the case with gold reported earlier. The adsorption rate was relatively slow overall, despite a rather fast initial adsorption. The apparent activation energy for the adsorption was measured to be 30.2 kJ/mol. It was believed that the kinetics of the silver adsorption on activated carbon was limited by a mixed control mechanism.

Jan 1, 1998

By III Kimball

For many years, Arizona merely shadowed EPA's environ-mental regulation of mining operations and waste. During this period, Arizona generally assisted EPA in the implementation of federal programs and enacted environmental laws patterned closely after their federal counterparts. A growing public concern about protecting water quality, particularly groundwater quality, eventually caused the State to depart from this pattern.

Jan 1, 1989

By K. A. Eriksson

Over 25 years ago, SKEGA pioneered the use of rubber linings in grinding mills and rubber is today widely accepted and almost standard material chosen for Regrind -and Secondary Mills. There are, however, only a few Primary Mills in operation with "solid" rubber linings because available materials and designs are often not cost efficient in comparison with steel alloys on a cost per ton basis and/or downtime for liner replacement would be too high.

Jan 1, 1988

By L. R. Duncan, W. H. McCracken

Magnesium is the eighth most abundant element in the earth's crust and the third most plentiful element in seawater. It is found in more than sixty minerals and in brines and seawater as a magnesium salt. One of the more important magnesium minerals is magnesium carbonate, MgCO3, with a theoretical maximum magnesia, MgO, content of 47.6%; this carbonate form represents the world's largest source of magnesia. The next most used sources for magnesia are magnesia-rich brines and seawater, where it occurs as magnesium sulfate and magnesium chloride. Other commercially important magnesium-bearing minerals are dolomite, CaMg(CO3)2, which serves the aggregate industry as well as the chemical industry; brucite, Mg(OH)2, which is used in the production of both caustic and sintered magnesia; olivine, ([MgFe]2-SiO2), which serves the refractory and heat storage industries; talc, (H2Mg3(SiO3)4), which serves several industries as a filler and as an ingredient in cosmetics; and serpentine, (H4Mg3SiO9). All these minerals are of commercial value because of the desirable characteristics given to them by magnesia and by its placement in the crystal structure. These minerals are the starting raw materials for a wide range of products. These include the production of magnesium metal, and several grades of magnesia used for the production of both sintered magnesia for refractory manufacture and lighter fired caustic magnesia, the latter is used in fluxes, fillers, insulation, cements, decolorants, fertilizers, chemicals, in the treatment of waste water including pH control, and in the removal of sulfur compounds from exhaust stacks. Magnesite is one of several major minerals possessing a dual character, i.e., non-metallic industrial mineral because of its principal end use as a calcined/dead burned raw material in non-metallic end products and magnesium metal. A large percentage of the magnesia used to produce magnesium metal comes from seawater1 brines. The total primary metal production worldwide is about 350 ktpy. This report reviews magnesite solely as an industrial non- metallic raw material. Magnesite for most refractory purposes must have a content of sintered magnesia above 85%, with a preference for those sources that have the potential for 90 to 98% magnesia values. Dolomite, used in large tonnages because of its physical properties in the construction industry, also has a use in the chemical and the refractory industry. In its calcined form, dolomite is used as a slag addition in the making of steel, and it is used in a high-purity calcine to precipitate magnesium hydroxide from brines or seawater. The precipitated magnesium hydroxide is intermediate for the production of deadburned clinker for refractories and the production of magnesium metal. In many cases the high-purity magnesium hydroxide obtained by precipitation with calcined dolomite, or in some cases limestone, has supplanted much of the world's production of magnesia from the natural mineral magnesite. This is especially true of the United States and in major industrialized areas of Europe and Japan. The best known of the minerals directly and widely exploited for its magnesia content is magnesite (MgCO3), one of the calcite group of rhombohedra1 carbonates, which includes calcite (CaCO3), siderite (Fe2CO3), and rhodocrosite (MnCO3), among others. The members of this group enter into a wide range of substitutional solid solutions when the positive ions have similar radii. The radii of magnesium and iron ions are within 6% of each other; hence magnesite and siderite form a complete series, of which the mineral breunnerite (ferroan magnesite) is a well known end member. The radii of calcium ion is 36% larger than that of magnesium ion and only limited substitution exists at each end of the MgC03-CaCO3 series. Dolomite is not a member of the calcite group, but it occurs when calcium and magnesium ions alternate in equal number in an ordered structure among carbonate ions. The result of this relationship is that calcite and dolomite are commonly found intermixed with magnesite. They occur, commonly, as identifiable crystal entities, which can be separated to a varying degree from the magnesite by benefication techniques. Magnesite, when pure, contains 47.8% MgO and 52.2% CO,. The pure mineral is found occasionally as transparent crystals resembling calcite. Magnesite principally contains variable amounts of the carbonates, oxides, and the silicates of iron, calcium, manganese, and aluminum. Although the genesis of natural magnesite deposits can be complex, it is distinguished in nature in two distinct physical forms, namely crystalline, (with a wide range of visible crystal sizes) and cryptocrystalline, sometimes referred to as amorphous, where the crystal size is not detectable to the eye and will range from 1 to 10 pm. The two types not only differ in crystal structure but in the sizes of the deposits and in modes of formations. The crystalline form has a Mohs hardness of 3.5 to 4.0. The color range is from white to black with shades of yellow, blue, red, or gray. The color is not a significant indicator of purity but in a given deposit an experienced person can often roughly grade the magnesite by observing color and crystallinity. Macrocrystalline deposits occur in relatively few, but generally large, deposits on the order of several million tons. The ore shows a marble-like crystal structure and belongs to the sedimentary or metasomatic groups of origin. The cryptocrystalline variety of magnesite frequently occurs in many small deposits, although there are exceptions. Cryptocrystalline magnesite is typically massive with no cleavage and is sometimes descriptively called bone magnesite. The fracture is usually conchoidal, with a hardness of 3.5 to 5.0. The color is mainly white, but it can have tints of yellow, orange, or buff. Accessory silaceous minerals such as serpentine, quartz, or chalcedony are usually present. Calcium minerals are normally absent or in low concentration, this contrasting with the macrocrystalline form where calcium minerals are almost invariably present and some- times in high concentration. The specific gravity of cryptocrystalline magnesite ranges between 2.90 and 3.00, whereas the value of pure crystalline magnesite is 3.02. In actuality, the normal specific

Jan 1, 1994

By E. P. Gustafson

Demand by precious metals explorationists for cost effective means to target subsurface geologic structures worthy of evaluation by means of drilling has led to the use of the predictor-correlator technique. To locate polarizable targets, the predictor-a localized inhomogeneity in electrical properties-is placed at various locations in the computer-simulated subsurface region, and correlated with the potential distribution observed in the field. The correlation coefficient obtained is assigned to the subsurface location. IP targets are extracted by calculating a second correlation section based on higher-frequency illumination and subtracting it from the first. The locations of correlation difference correspond to the locations of polarizable mineralization.

Jan 1, 1989

By K. H. Bettles

The Goldstrike Mine is in Eureka County, Nevada, approximately 23 miles northwest of the town of Carlin. The mine is at the north end of the Tuscarora Spur of the Tuscarora mountains at elevations of 5,300 feet to 5,950 feet. It is part of the Bluestar/Goldstrike subdistrict of the Carlin Trend of sediment hosted, disseminated gold deposits. The Carlin trend comprises over 15 major gold deposits with production, reserves and geological resources of approximately 50 million ounces of gold (Fig. 1). In the Bluestar/Goldstrike subdistrict, there are four major deposits; Post, Betze, Bluestar, and Genesis, plus numerous smaller deposits (Fig. 2). Betze is located on property controlled by Barrick Goldstrike Mines; Bluestar and Genesis on Newmont Gold property; and Post on both Barrick and Newmont property.

Jan 1, 1989

By M. J. Meulendyke

Within the mineral processing Industry. a range of grinding conditions exists which include semi-autogenous grinding (SAG), rod milling. and conventional ball milling. Each of these mill environments presents a unique environment for grinding media, requiring the application of specific physical and chemical properties for optimum grinding media performance. The environments are characterized by varying degrees and combinations of abrasive, corrosive, and impact wear. An extensive test program has been conducted to determine the extent wear rates vary between these different applications. Test results are related to production results, ball size, and mill operating conditions.

Jan 1, 1989

By M. Murphy

The multibillion dollar cost of cleaning up thousands of environmentally contaminated sites across the country created by decades of inadequate disposal of hazardous waste has led to a search for "deep-pocket" sources of cleanup funds. The resul1:: is the emergence of a growing body of law that could have a profound impact on commercial financing transactions and other investment activities when present or past owners of contaminated property are involved.

Jan 1, 1988

By T. D. Maki

FMC Gold's Paradise Peak Mine, located 13 km (8 miles) south of Gabbs, Nevada, became operational in April of 1986. Shortly thereafter heap leach testwork began on sub mill grade ore. Presently, testwork continues in combination with production size heaps that account for approximately 6% of total gold production at Paradise Peak.

Jan 1, 1989