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By R. D. Hagni

Introduction The ore deposits of the Viburnum Trend in the Southeast Missouri Lead District constitute the world's largest lead producing district. Six or seven mines currently are in production in the Viburnum Trend. The principal minerals are few in number, simple in composition, and form simple superposition and replacement textures. The economically valuable minerals in the typical lead-zinc ores of the Viburnum Trend are the sulfides, galena, sphalerite, chalcopyrite, and siegenite. The most abundant gangue minerals are dolomite, marcasite, pyrite, calcite, quartz, and dickite. Four concentrates, lead, zinc, copper, and cobalt are produced from the ores, and they contain an interesting suite of trace elements, of which silver and cadmium are recovered at the smelter and refinery.1 Emphasis is given here to the variations in directions of ore fluid flow based upon interpretations of the positions of orebodies due to the ore fluids being impeded by Precambrian basement relief and upon patterns of ore mineral/metal zoning. The Precambrian basement relief and Lamotte Formation pinchout has exerted an important control upon the specific locations of orebodies in the Southeast Missouri Lead District. The locations of some orebodies along one side of Precambrian knobs or ridges strongly suggests that they are positioned on the side from which the ore fluids flowed. Other orebodies are located in positions that suggest that they formed where the ore fluids passed around or between Precambrian knobs. Patterns of mineral and/or metal zoning have been used in many mineral districts to interpret the directions of ore fluid flow. The mineral/metal zoning pattern correlates closely with that of the mineral paragenesis. The earliest deposited minerals tend to be deposited nearest to the source, the later deposited minerals next closest to the source, and the latest deposited minerals furthest from the source. In the Southeast Missouri Lead District, it is interpreted here that the early-deposited copper-rich ores are closest to the source of the ore fluids, the subsequently-deposited zinc-rich ores are next closest to the source, and iron-rich ores are furthest from the source.

Jan 1, 2006

By James H. Boettcher

Inflation and increasingly complex tax laws in the United States combined with rising national ism overseas have made the structuring of viable financing plans which meet the requirements of the various parties involved equally important in determining project success as the design of an efficient mining plan or selection of the optimal size equipment. Whereas the use of computerized analysis to design and evaluate the technical parameters of resource ventures is well accepted, computerized evaluation of the investment criteria (qualitative and quantitative) of the various parties to the project (mining firm, financial institution, product buyer, host government, etc.) is a more recent application which can make a material impact in strengthening project economics. By identifying the objectives of the parties and critically evaluating and testing them with the technical and economic/financial constraints on the project the sponsors can design a financial structure tailored to the attributes of the deposit. A myriad of new financial techniques have evolved for financing natural resource ventures, from "Commodity-backed bonds" to tax based limited partner- ships, but in every case the structure of the financing mechanism depends on carefully matching the parameters of the financial structure with the economic characteristics of the project and the objectives of its sponsors. This presentation outlines a methodlogy for computer based project evaluation in the design and structuring of stronger financial plans for natural resource ventures.

Jan 1, 1982

By D. Nilsson, B. Aaro

To find out if there are any potential savings in "in-the-wall haulage for open pits," the Swedish companies ASEA and Kiruna-Truck in 1984 gave us financial support to study this solution in more detail. In the study, the use of in-the-wall haulage was studied for three hypothetical open pits with different sizes and shapes. Annual ore production rates also varied. A part of this thorough study was published by the Swedish Mining Research Foundation (see Report F 8444, "The Use of Electric Trucks and In-the-Wall Haulage in Open-Pit Mining," in Swedish with an English Summary). Although in-the-wall haulage is of interest in some open pits, the local terrain is important for profitability. The authors do not think in-the-wall haulage is of any major interest for the mining industry. The following is a summary of some of the author's findings. To place the haul road in the wall is not of interest in open pits with declining metal contents in the bedrock. In such mines, the volumes from the haul roads will not yield revenues. The cost per m3 for an underground haul road is much higher than for a haul road in the pit. This means that it is only of interest to place the haul road in the wall when mining gets deeper than about 50 m (165 ft). In-the-wall haul roads will reduce flexibility in the pit, and it will make necessary the use of smaller equipment with lower productivity and higher costs per ton. As Hustrulid, et al. show (Table 4), excavation savings of material hauled are very low, $0.10 to $0.25/t (0.09 to 0.27 per st). The extra operating cost, due to lower productive equipment, will normally be much higher and thus destroy the whole idea with in-the-wall haulage. If electric trucks are profitable for in-the-wall haulage, it is normally also profitable to use electric trolley assist for trucks on the haul road in the pit. But the profitability of using electric power is different in different countries, and depends on the relation between the cost for electric power and diesel fuel. In the US, diesel fuel is inexpensive compared with electricity, but in a country like Sweden, diesel fuel is much more expensive. In most open pits, the trucks have to move a considerable distange from the loader until they reach a final haul road in the wall. A trolley line along such temporary haul roads will be exposed to flyrock. It is normally less expensive to perform rock support from the open pit than from underground ramps in the wall. In Fig. 8, Hustrulid et al. gives the impression that the underground haul roads will be very close to each other. This is seldom the case. Figure 1 shows a haul road in the wall. The haul road passes through each cross-section only once. Arranging a reliable dewatering and rock support system from only one underground ramp is probably impossible. Many more drifts are therefore necessary. In our report, we also studied what would happen if the final pit slope is increased by 5°, using underground drifts in the wall. Our conclusion was that, the extra cost for drifts, rock support, pumps, etc. destroys the whole idea, and that it was better to accept a higher stripping ratio with the haul road in the pit and to use conventional low-cost open pit equipment. Finally, we think it is would be difficult to try to reduce the safety factor when determining the slope angle by moving the haul road in the wall. Minimizing the risk for pit wall collapse is also important with in-the-wall haulage, primarily because men and equipment will be working in the pit, but also to guarantee many accesses between the pit and the haul road in the wall.

Jan 1, 1990

By M. Albrecht

I have been following the discussions on coal degradation started by the article "Degradation process in coal slurry pipe-lines," by M.G. Ayat and B.C. Scott. I agree with Mr. Dasher's comments in January that the title of the article was perhaps misleading as the information appeared to pertain more to processing the coal than to transportation. Specifically, the authors' conclude that the cyclone had an important impact on the degradation. Dr. Ferri's comments on degradation in other physical processes was very correct, handling of coal in a turbulent manner has an impact on the size distribution of the coal. This is one reason that when I design a coal system (mine to rail car) I try to minimize the total number of transfer and processing steps consistent with achieving the desired product specifications. Another area, mentioned in passing in the article, the impact of natural causes and the mining step, is of particular interest to me. A mine in West Virginia had a variation of 50% in the -6.35 mm (-0.25 in.) raw coal depending upon the direction of mining. Since it was a single unit mine, this significantly impacted the final product. Two articles on mining's impact on size distribution are: • Stefanko, R., Chopra, 1K., and Ramani, R.V., 1974, "Mining influence on size consistency and washability characteristics of coal," SME Fall Meeting, September; and • Albrecht, M., and Olsen, D., 1980, "Mining methods impact on coal preparation plant design," 1st International Mine Planning and Development Symposium, September. Both articles presented information on variation of size distribution and cleaning plant performance on a coal depending on the mining method used. Both articles considered a single coal seam at one mining operation. Stefanko's article considered continuous and conventional mining of coal in Pennsylvania. Albrecht's article considered surface, underground continuous, and underground hydraulic mining of a coal in British Columbia. Both concluded that mining methods do impact coal characteristics. I agree with Mr. Ayat's comments in reply to Dr. Ferri's discussion, "... degradation is a multi-faceted phenomenon involving many natural and man-made factors." As efforts are being made to produce a higher quality (lower sulfur) coal, this topic becomes even more important

Jan 1, 1989

By S. Serata

Background Results of the above study in the August 1991 issue of Mining Engineering offer valuable lessons in the solution of cutter-roof problems. The original study plan was initiated by the discussion authors to solve the problems using the "stress control method" of mining (Serata 1976,1982; Serata, Carr and Martin, 1984; Serata and Gardner, 1986; Serata, Gardner and Preston, 1986; Serata, Gardner and Shrinivasan,1986; Serata and Kikuchi, 1986; Serata, Preston and Galagoda, 1987) However, the plan and the planner were changed to the arrangement reported in the paper. The change was considered reasonable at the time due to the mine engineers' uncertainties about the stress control method. Consequently, the basic principle of the study was shifted from the original stress control method to the method described in the paper, which will be called the "yield pillar method" for the purposes of this discussion. The paper convinces the reader that the yield pillar method fails to solve the cutter-roof problems. This doesn't mean that the stress control method also fails. Actually the contrary is true, as discussed below. Limitation of the yield pillar method The paper illustrates clearly how poorly the yield pillar method performs in solving the problem. The reason for this failure is the lack of the protective stress envelope needed to stabilize the cutter roof. Unfortunately, the protective envelope cannot be formed properly without utilizing the stress control method of mining. Changing the pillar size does not make much difference in the roof stability.

Jan 1, 1992

By D. Nilsson, B. Aaro

To find out if there are any potential savings in "in-the-wall haulage for open pits," the Swedish companies ASEA and Kiruna-Truck in 1984 gave us financial support to study this solution in more detail. In the study, the use of in-the-wall haulage was studied for three hypothetical open pits with different sizes and shapes. Annual ore production rates also varied. A part of this thor¬ough study was published by the Swedish Mining Research Foundation (see Report F 8444, "The Use of Electric Trucks and In-the-Wall Haulage in Open-Pit Mining," in Swedish with an English Summary). Although in-the-wall haulage is of interest in some open pits, the local terrain is important for profitability. The authors do not think in-the-wall haulage is of any major interest for the mining industry. The following is a summary of some of the author's findings. To place the haul road in the wall is not of interest in open pits with declining metal contents in the bedrock. In such mines, the volumes from the haul roads will not yield reve¬nues. The cost per m3 for an underground haul road is much higher than for a haul road in the pit. This means that it is only of interest to place the haul road in the wall when mining gets deeper than about 50 m (165 ft). In-the-wall haul roads will reduce flexibility in the pit, and it will make necessary the use of smaller equipment with lower productivity and higher costs per ton. As Hustrulid, et al. show (Table 4), excavation savings of material hauled are very low, $0.10 to $0.25/t (0.09 to 0.27 per st). The extra operating cost, due to lower productive equip¬ment, will normally be much higher and thus destroy the whole idea with in-the-wall haulage. If electric trucks are profitable for in-the-wall haulage, it is normally also profitable to use electric trolley assist for trucks on the haul road in the pit. But the profitability of using electric power is different in different countries, and depends on the relation between the cost for electric power and diesel fuel. In the US, diesel fuel is inexpensive compared with electricity, but in a country like Sweden, diesel fuel is much more expensive. In most open pits, the trucks have to move a considerable distange from the loader until they reach a final haul road in the wall. A trolley line along such temporary haul roads will be exposed to flyrock. It is normally less expensive to perform rock support from the open pit than from underground ramps in the wall. In Fig. 8, Hustrulid et al. gives the impression that the underground haul roads will be very close to each other. This is seldom the case. Figure 1 shows a haul road in the wall. The haul road passes through each cross-section only once. Ar¬ranging a reliable dewatering and rock support system from only one underground ramp is probably impossible. Many more drifts are therefore necessary. In our report, we also studied what would happen if the final pit slope is increased by 5°, using underground drifts in the wall. Our conclusion was that the extra cost for drifts, rock support, pumps, etc. destroys the whole idea, and that it was better to accept a higher stripping ratio with the haul road in the pit and to use conventional low-cost open pit equipment. Finally, we think it is would be difficult to try to reduce the safety factor when determining the slope angle by moving the haul road in the wall. Minimizing the risk for pit wall collapse is also important with in-the-wall haulage, primarily because men and equipment will be working in the pit, but also to guarantee many accesses between the pit and the haul road in the wall.

Jan 1, 1989

By J. R. Sturgul

Thompson and Adler (1988) present a new simulator for use in designing conveyor belt systems for underground mines. Their model uses a large Fortran program to do a Monte Carlo simulation. Their approach is similar to that done by other researchers, such as Stanford (1965) and Newhart (1977). All Fortran based simulation models use an imaginary clock that is advanced a time increment. At each advance of time increment the program has to determine what is to take place in the system. Since many things can take place at different places in the system, this approach requires that a large amount of logic is needed at each step of the program. This, then, necessitates a great amount of jumping between different portions of the program. In fact, the Fortran program, CMBCS, takes up to 15 to 20 minutes to run on the mainframe and several hours for the PC. Such execution times are typical of Fortran based simulation studies. There have been other examples of Fortran based simulation studies in mining. Cross and Williamson (1969) presented a model of a working open-pit copper mine. Fosinato (1982) presented a model of the bucketwheel excavators for the large open-pit mines in Southern Victoria in Australia. These excavators discharge into a complex conveyor system, which can be thought of as being similar, but more complicated than the one modeled by CMBCS. All previous simulation studies using Fortran as the source code have several things in common: they all are very lengthy, often being of the order of 10,000 lines of code and take a great deal of CPU time to execute. The program of Fosinato takes more than 4.5 hours to run on a large mainframe. Also not mentioned by Thompson and Adler is the actual time it took to write the Fortran code. This, too, can be considerable, as Fosinato's program took several man years to write and de-bug. In order to change it, as the parameters change or as the mine increases, further large amount of time is needed. An alternate to such Fortran based programs is provided by using a special computer language known as General Purpose Simulation System (GPSS). This language is designed to solve a wide variety of discrete event simulation problems and appears to be particularly suited for mining situations as shown by Sturgul and Yi (1 987). Although GPSS was introduced as far back as 1961, it has not been used in mining to any great extent. However, the original form was quite crude and inefficient compared to modem versions. Perhaps with the increasing popularity and availability of these recent versions, this will change. Two such versions are GPSS/H for the mainframe and GPSS/PC for the personal computer. GPSS/H allows for interactive usage, such as allowing input data to be entered via a menu. GPSS/PC has excellent graphical capabilities and one can actually view the simulation in progress. In addition, one can halt the simulation and change values of various parameters. If a particular portion of the simulation is in progress via a concept known as "microwindows." These new versions have greatly increased the potential for using the language for mining situations. Mining applications of GPSS have been given by Sturgul and Yi (1987), Sturgul and Harrison (1987), Sturgul (1987), and Sturgul and Singhal (1988). Schriber (1987) presents a summary of the recent advances in GPSS.

Jan 1, 1987

By George F. McKeregahan

Phosphate production of the Central Florida Phosphate District has been predominantly from the unique and lucrative Bone Valley Formation, with local and minor amounts coming from the underlying Hawthorn Formation. As the Bone Valley is depleted, replacement production will move progressively into areas of total Hawthorn Production. Significant changes in the characteristics and make-up of the deposit dictate required modifications in mining and beneficiation. Failure to tailor technology to fit the deposits of the future will result in vast "reserve" losses. In the Southern Extension of the Central District - two distinct "ore types" exist. These are presently referred to as the upper (high grade) zone and the lower (low grade) zone. The upper zone is somewhat similar to the Bone Valley Formation and will respond favorably to present beneficiacion practices. The lower zone is strikingly different and its response to conventional beneficiation is not acceptable. The major differences are in grain size distribution and mineralogical make-up of these grains. The Hawthorn differs from the Bone Valley by exhibiting lower grades, lower tonnage yield per acre foot, lower pebble percent of product, higher mining ratios, higher slimes (-150 mesh) content and deeper mining depths. Physical characteristic changes include more frequent occurrences of competent bands or lenses of "limerock" which are submarginal in value.

Jan 1, 1983

By G. Meiers, M. O’Kane

Dry cover systems as a closure option for management and decommissioning of waste rock and tailings is a common prevention and control technique used at numerous sites around the world. Numerous site-specific factors control the long-term performance of dry cover systems. Site-specific climate conditions are one of the key factors to consider when determining the objectives of a dry cover system because these conditions will significantly influence long-term performance. There are advantages and disadvantages of characterizing the site on an annual basis as generally a “dry” site (annual potential evaporation greatly exceeds annual precipitation), or generally a “wet” site (annual precipitation meets or exceeds annual potential evaporation). The field performance monitoring data and discussion presented herein illustrates the caution that is required when using these “annual” criteria for characterizing site climate conditions. Numerous sites exist where precipitation exceeds potential evaporation on an annual basis; however, the site typically experiences hot, dry summer months where evaporation greatly exceeds rainfall. These dry summer conditions have a strong influence on long-term performance. In contrast, sites where potential evaporation exceeds precipitation on annual basis, typically experience periods of the year when a moisture surplus occurs, which also impacts on long-term performance. This paper uses soil-atmosphere cover design modelling and field performance monitoring from sites in Canada, Australia, and the United States to demonstrate the importance of considering site-specific climate conditions when predicting long-term cover system performance.

Jan 1, 2003

By A. E. Hall

The authors are to be commended for their attempt to provide a convenient piece of software for preliminary drum hoist specification. This field has received little attention in recent years and is eminently suited to calculation by computer. Unfortunately, there appears to be an error in the calculation of the required rope diameter. The 66-mm (2 5/8-in.) rope selected is oversized for this application. Wire rope weight and breaking strength are related to the rope diameter. For a round strand rope

Jan 1, 1988

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 theirs 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 m (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 geologi¬cal 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 experimen-

Jan 1, 1991

By F. C. Bond

History The breaking and shaping of rock was one of man's earliest occupations. In the Paleolithic Age long before the dawn of history, arrow¬smiths and the makers of stone axes, hammers, knives, scrapers, spears, and borers were highly respected members of society. In early historical times stones for building blocks, roads, and city walls were shaped by slaves and convicts, who also did most of the mining. However, great artists erected beautiful stone sculptures, while gifted architects planned imposing temples and monuments. Until well into the 19th century nearly all rock was broken laboriously by hand. The small rock required by John MacAdam for his macadamized roads in England in the 1820s was produced by women and boys seated alongside the roadside with hand hammers and legs wrapped with rags. Eli Whitney Blake, a nephew of the Eli Whitney who invented the cotton gin, developed the first successful jaw crusher before 1870. The gyratory conical crusher soon followed. Comparative tests established its large capacity advantage over the jaw crusher, as well as its greater cost for a given feed size. Both types have been in use for more than 100 years. Crushing rolls appeared before 1900. Thomas A. Edison made very large diameter rolls which were excessively long; they failed because of shaft deflection. Various types of disk crushers and edge runner rolls appeared about this time. The older methods of reducing rock were adaptations of other processes. The stamp battery of dropping weights effected crushing by simulating heavy hammer blows; the much earlier arrastre, in which heavy stones were dragged in a circular path over the ore by animal power, came from the prehistoric method of grinding grain between two rubbing stones, while the jaw crusher was adapted from simple squeezing devices. But the tumbling grinding mill was not just an adaptation; it was an invention, because it required thinking on a somewhat higher order-there was no prototype. Its nearest antecedents were probably the small closed tumbling drums used in England more than a century ago for cleaning and polishing small iron castings. The date of the first tumbling mill actually used to grind rock is unknown, but it was later than the American Civil War (1861¬1865). It was almost certainly a closed or batch mill in which rock was placed and rotated until it reached the desired particle size. It could have been operated either wet or dry. The first published refer¬ence to such a batch mill was one introduced by Alsing in England (1870) for the grinding of calcined flints for pottery work.21 There are several rather indefinite reports of grinding mills in the early 1890s, including an overflow ball mill in the Helena and Livingstone reduction plant in Montana which may have been the first of its kind. 11 Many of the first mills, which were called tube mills, used hard rock both as grinding media and as mill lining. The rock used was preferably stone from the Normandy and Danish beaches, when it could be imported. This remarkable siliceous stone was already widely used for grist mills throughout America, and its resistance to wear was greatly respected. The decade of the 1890s saw the development of tumbling mills with continuous feed and discharge and their extension into different industrial uses. By 1895 some experience had been accumulated. Iron grinding balls were being tested and the proper speed of rotation was being determined. The Clark Patent tube mill was featured in an E. P. Allis bulletin of 1890, which may have been the first published description of a tube mill. More than 1,000 Gates tube mills had been built by Allis-Chalmers before 1913. Many of these were used in gold mining, espe¬cially in South Africa. The 5 x 22-ft size was particularly favored for grinding portland cement; the use of tumbling mills in the manufac¬ture of cement began about 1900. A great deal of attention was paid to the mill lining. Metal was then relatively expensive, and the general approach was to trap some of the rock grinding media into mill lining pockets. This rock would then absorb the wear and protect the metal lining. In the first ten years of the 20th century there were several different types of pocketed liners, with different manufacturers advancing the superior claims of their patented arrangements. The Osborne liner, developed in South Africa, was probably the most successful. 21 Another item which attracted much attention was size classifica¬tion within the mill and in ancillary equipment attached to the mill and rotating with it. The Krupp type, with interior screens protected inside the mill lining, was developed quite early in Germany, possibly before 1890. The Dorr reciprocating rake classifier (1907) had not yet been invented, and many strange and impractical screening and classifying devices were proposed. In these unsatisfactory machines the two separate processes of size reduction and classification were combined into one operation. It was many years before recognition came that a machine is most efficient when it is designed for one specific purpose. There was much industrial wastage before the opera¬tions of grinding and classification were finally separated. After 1900 the grinding of portland cement raw material and of cement clinker required large numbers of tumbling mills. Most of the raw material was then ground dry. This was also the heyday of gold mining. The old stamp mills that were used in great numbers for grinding gold ore did not grind sufficiently fine to liberate all of the gold, and the new tube mills were installed following the stamps. After 1910 larger diameter tum¬bling mills with larger grinding media were developed. These could receive the finely crushed ore directly, and the inefficient stamp batter¬ies were gradually eliminated. The Rand in South Africa was the greatest gold producer, treating immense quantities of rather low-grade but consistently free milling gold ore. The first tumbling mills, or tube mills, went into operation there in 1904. They were so successful that within a few years no new stamp batteries were installed in the district, even though old ones continued to pound away until after 1950.3 The early tube mills on the Rand all employed Normandy or Danish pebbles, which had to be imported at considerable expense. Their reported wear was as low as 4 lb per ton ground. 4 Many of the mills were lined with the same tough Danish stone cemented into place, while others used the pocketed steel Osborne liner. It was in 1907 on the Rand that an important test was made using hard native ore for grinding media in place of the expensive imported pebbles. 3 This ore, called banket, did not wear as well as the renowned Danish pebbles, but the cost per ton of grinding was definitely reduced. Many of the tube mills on the Rand were soon grinding with native ore. This was the beginning of the development now called autogenous grinding, in which the ore grinds itself. This is treated under a separate heading. See Subsection 3C, Chapter 4. Gold mining was important in America also, and its grinding history follows that of the Rand. Danish pebbles were replaced by native ore in Santa Gertrudis, Mexico, in 1913, 4 and in Consolidated Gold Fields, Nevada, in 1914. 5 Other properties followed suit. However, it gradually became apparent that the capacity of a given mill could be almost doubled if rock grinding media were re¬placed by cast iron or steel grinding balls. In order to increase plant grinding capacity many rock media mills were converted to iron grind¬ing media in the second decade of the present century. In some mills the motor size was doubled; other mills were cut in two and another motor was provided for the second half. Grinding mills began to assume a more modern appearance. Crushing rolls were formerly much used following crushing in jaw or gyratory crushers and preceding grinding in ball mills. How¬ever, the roll surfaces wore rapidly, and skilled maintenance was required to obtain even wear. Rod mills could take feed of the same crusher product size and reduce it finer. The first rod mill was con¬structed by Mine and Smelter Supply Co. and was tested in Canada

Jan 1, 1985

By J. L. Weeks

Discussion by J.L. Weeks Discussion I wish to comment on Kissell and Sacks's paper "Inaccuracy of area sampling for measuring the dust exposure of mining machine operators in coal mines." It is well established in industrial hygiene that area sampling is not an accurate measure of personal exposure. To measure personal exposure, one must use a personal sampler. Kissell and Sacks, using data from coal mining, reached the same conclusion. This is the right answer but it is the wrong question. This issue arises because the machine-mounted respirable dust continuous monitor (MMRDCM) is being considered to monitor coal-miners' exposure to respirable dust. This device may be mounted on a continuous mining machine, and it would take a modified area sample. I refer to it as "modified" because a mining machine is not really a fixed location, as is the typical area sample. The concern with whether this machine would provide an accurate estimate of the machine-operator's exposure, however, overlooks important features of the MMRDCM. Specifically, it displays dust concentration as it occurs and at the place where it is needed, not 10 days later, as is the case with the current system. Used in this way, it allows for timely intervention to control exposure to dust and it leads to greater first-hand knowledge of specific conditions and mining practices that result in excess exposure. If one knows the dust level is excessive at the time the dust is being generated, it is possible to use known and effective methods of dust control to reduce exposure. By measuring again, one can determine whether control methods have worked. This is a generic procedure followed in any industrial process: measure, adjust and measure again. It leads to more effective and more efficient dust control and is preferred to taking a sample, waiting 10 days for the results and then trying to apply that information to a situation that no longer exists. The MMRDCM also records the dust level automatically, and it runs continuously, making it harder to tamper with. Dust monitoring in underground coal mines has been corrupted by some operators having submitted fraudulent samples (and being convicted on criminal charges because of it), weakening the credibility of the current system. The MMRDCM is, in part, a response to this corruption, and an effort to focus attention on dust levels and dust controls rather than on the perpetrators of fraud. Use of the MMRDCM is an evolutionary change from the current system. With the current system, one takes a personal sample of exposure of the "designated occupation" (DO, defined at 30 CFR 70.2 (f)), the occupation regularly exposed to the highest level of dust on a section. Typically, on sections using continuous mining machines, this is the mining-machine operator. The rationale behind DO sampling rests on some of the findings that Kissell and Sacks report. They reviewed several studies that showed that dust concentration varies with location on a mining section and that there is a concentration gradient with the highest concentration occurring closest to the face. The logic behind DO sampling strategy is that if the dust concentration at the place where it is highest is below the exposure limit, then concentration elsewhere on the section is also below the limit. In this way, one will have achieved the objective or keeping dust concentration below the exposure limit for all miners on the section by measuring the exposure of only one miner. The

Jan 1, 2003

By Mike Wardell-Johnson, Chris Campbell-Hicks, Craig Bitmead

Faster, simpler, cheaper and more accurate cyanide ex­tractable gold procedures are now available for grade control assay. Cyanide 'bulk leach extractable gold' or BLEG assays have previously been seen as desirable, due to the large sample size that can be treated, but not in common use due to the 24hr total assay time required. Sample sizes from 50g to several kilos can now be accurately assayed in less than 4 hours total time with the larger sample size increasing accuracy, without loss of precision, while greatly reducing the 'nugget effect' or coarse gold bias. Typical assay size is 200g, equivalent to 4 x 50g Fire Assays or 10 x 20g Aqua Regia Assays. Ongoing modifications and refinement of the procedures and reagents used also indicate applicability to the treatment of coarse gravity gold where substantial savings in both capital and operating costs for these circuits are possible. The successful use of this cyanide assay at the Bannockburn and Windine Well operations are cited to-gether with a compre­hensive comparative testwork program by Newcrest. Testwork at Gidgee and Lynas Gold operations is currently being initiated. Other operations using the method include Red Dome and Kidston; with Placer(Granny Smith) and PMA's Palm Springs operations to commence shortly.

Jan 1, 1995

By A. W. Khair

This paper presents an analysis of respirabie dust generation due to the action of a continuous miner. Underground coal cutting by a drum-type continuous miner was simulated in the laboratory using a specially designed unique automated rotary coal cutting simulator (ARCCS). An analysis of -400 mesh particles was done on gravity collected material. Samples of entrained dust were collected through the use of cascade impactors. Both samples were analyzed for size distribution and particle shape. Using the simulated coal cutter permitted variation of operating and in-situ parameters to provide information on the effects on generation of -400 mesh particles. In these tests coal blocks with an approximate dimension of 18 in. x 15 in. x 6 in. (45.7 cm x 38.1 cm x 15.2 cm) were first subjected to confining pressures, equivalent to in-situ conditions. Such blocks were then cut by the cutting head of the ARCCS, thus simulating the action of the continuous miner in underground coal mining. During the tests under a particular set of in-situ and operating parameters a number of other parameters such as 1) penetration of bit into coal, 2) penetration resistance (thrust and cutting pressures), 3) rotating velocity of cutting head, and 4) acoustic emission activity in the coal block were monitored. After each cutting cycle the fractured surfaces were photographed and a velocity survey was conducted by using a sonic technique. At the end of each experiment the cutting paths of the bits in coal were photographed using an optical microscope with an attached camera.

Jan 1, 1985

By R. H. Goodman

The development of techniques for separating minerals based on differences in their specific gravities is reviewed and their present applications in process technology outlined. Special emphasis is placed on the development, in Australia, of new models of the spiral concentrator. The mechanics of separation of spiral separators are stated, including the effects of operating parameters, Some of the traditional areas of spiral application are described, together with examples of new applications such as treatment of fine alluvials and tailings, separation of liberated minerals at coarse sizes and fine coal recovery.

Jan 1, 1984

By R. J. Sweigard

The effects of postmining land use planning are addressed in this paper. These effects may include savings in reclamation costs and increased value of reclaimed land. However, this may result in greater planning costs and possible delays in production. Therefore, an economic analysis was undertaken in an effort to determine the importance of each potential result and to quantify the overall economic impact of postmining land use planning. A discounted cash flow analysis was performed using a surface mine cost model to analyze the sensitivity of selling price to changes in various postmining land use planning parameters. As a result of this analysis it was determined that the profitability of a project is most sensitive to delays in production, particularly, if the delay occurs after major investments have been made. Many methods have been proposed for evaluating postmining land use alternatives (Sweigard and Ramani, 1981). Both quantitative methods, such as benefit-cost analysis (Brock and Brooks, 1968), and nonquantitative methods have been proposed. The application of one specific analytical method is pursued in this paper.

Jan 1, 1985

By Richard J. Fantel

The U.S. Bureau of Mines performed engineering and economic analyses on 130 domestic phosphate mines and deposits, all containing demonstrated resources, of which 73 are in the Southeastern states of Florida and North Carolina, to determine the domestic availability of phosphate rock. Of the total domestic resources studied (6.4 billion tons of recoverable rock) over 60 percent is found in these two Southeastern states. At an estimated cost of $30 per ton, in January 1981 dollars, approximately 1 billion tons of rock is potentially recoverable from Southeastern deposits, increasing to 3 billion tons at $45. In addition, there is nearly 6 billion tons of recoverable rock at the inferred level (some presently unmineable) in Florida and Georgia.

Jan 1, 1983

By Brian P. Canning

Ion exchange resin processes are receiving renewed attention in the mining industry for recovery;, of metals from solution. This interest has been inspired by the development of continuous countercurrent ion; exchange systems (CCIX) which are of much lower cost than the original, fixed bed contactor processes. With the countercurrent systems and new developments to process or upgrade the eluate product, ion exchange can now compete successfully with solvent extraction for treatment of uranium ores. The low concentration of U308 obtained in the CCIX strong eluate is generally upgraded by solvent extraction methods in sulfuric acid systems. Resin loadings of up to 120 grams per liter of U308, however, have been produced experimentally by recycling of eluate (HIM-IX process) and further developments may be expected in the near future. New developments in resin technology are bringing forth products of greater strength, loading capacity, selectivity and density. Densities up, to 1.35 spgr have been claimed by some suppliers. The process is not a new one and has taken years to advance to the present state of development. The earliest system to gain widespread acceptance was the resin-in-pulp system applied to uranium recovery. Ion exchange, however, later lost acceptance in favor of solvent extraction in the 1960's due to problems in resin degradation, resin contamination and resin-pulp separation, in addition to the development of improved solvent extraction reagents and techniques. The resin-in-pulp system continues to have application on high clay ores and has advanced in design. A number of other continuous ion exchange systems were also tried but never gained continued acceptance. In the current resurgence period the CCIX systems to gain acceptance were the Utah and Chem-Sep (Higgins Loop) systems. These processes are relatively simple in concept although require sophisticated controls. Fixed bed columns have also been designed into multistage systems in countercurrent operation but are not discussed in this paper. These systems remain competitive today for many applications. The most recent advancements however have been in the multistage contactors from,, which three are in current acceptance. These are the Himsley, Porter, and NIM systems. All industrial applications of these systems are relatively new, although pilot plant and in-plant development has been in progress over the last decade. Fur36ther developments should be expected in the near future.

Jan 1, 1979

By Stephen L. Bolivar

The Prairie Creek intrusion, near Murfreesboro, Arkansas, is the site of the only commercial diamond mine in North America. Three major lithologies are associated with the intrusion, a hypabyssal peridotite, a volcanic breccia in .which the majority of diamonds occur, and associated tuff. Previous studies, including mining history, geological, geophysical, petrological and whole-rock and mineral chemistry are reviewed. The recent suggestion, by other workers, that this intrusion may be more correctly called a lamproite is examined and supported. This has serious implications for diamond exploration in Arkansas and elsewhere in the U.S.

Jan 1, 1984