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By L. J. E. Hofer, L. F. Zeiler, D. A. Giardino

The Jones Trickett Ratio (JTR) is generally used in investigations of fires and explosions for the determination of the composition of the fuel. This paper takes a close look at the JTR and finds it unreliable and ambiguous for the determination of fuel composition. A new index is presented, termed the (H/C) Index which uniquely determines the fuel composition. The (H/C) Index is simple to use requiring only the results of a single gas analysis of the combustion products. Experimental data testing the (H/C) Index for methane, propane, ethylene and coal dust are given, showing that the Index unambiguously determines the composition of the fuel. The (H/C) Index is also applied to the gas analysis data of several mine fires, illustrating how the Index changes with the progression of the mine fire.

Jan 1, 1993

By J. H. Fewings

The rapidly escalating capital costs for new projects, or extension of existing projects, has placed greater emphasis on maximum use of existing resources. Thus, when increased grinding capacity was required to achieve metal production from lead-zinc ores a number of alternate approaches were studied. The most cost effective way of achieving the required 5% increase in grinding circuit capacity was to install upgraded instrumentation and computer control of grinding - Fewings, Wickham and Wiseman (1). Once the computer installation cost had been borne by the financial returns from that project, other projects involving computer control were economically viable on an Incremental cost basis. These Included centralised control of tailings thickening and return circuit water distribution, centralised control of the crushing plant and use of the computer as a data processor for an on-stream analysis system for flotation slurries.

Jan 1, 1979

By Donald D. Haas

Contract mining has a definite place in the mining industry, both in the United States and around the world. A mine owner should evaluate the utilization of contract mining in the devel­opment and production of mining properties. This evaluation of contract mining should be performed at various times in the life of a mining prope1iy from the development and initial operation of a new mining project, periodically throughout the mine's life, and at mine closure. Many factors must be considered in ana­lyzing and evaluating the potential of contract mining and those factors are both quantitative and qualitative. Quantitative factors are those factors for which costs or measures can be determined or estimated objectively; and qualitative factors are those factors for which costs or measures must be determined subjectively and do not lend themselves to objective determination or estimation. Some of the qualitative aspects may provide only yes, maybe, or no answers while others will provide a tendency or partiality toward one answer or the other without a quantified cost. The major quantitative aspects include operating and capital costs, monitoring costs, capital preservation, acquisition and disposal of new and used equipment, direct payroll and administrative costs, recruiting and training costs, and employment and termi­nation of staff and hourly personnel. The major qualitative areas include corporate philosophy, risks associated with capital and operating costs, financing the project, operating control, operat­ing flexibility, recruiting and training of personnel, employment and termination, administrative concerns, and owner/contractor relationship.

Jan 1, 1998

By D. A. Dahlstrom

Crushing and grinding consumes about 1.3% of the nations produced electrical energy.( Furthermore, only one in the first ten highest consumers comes from an industry other than mineral processing. This is the grinding of wheat to make flour. The other nine are all in the metallurgical industry and are easily the largest consumers.

Jan 1, 1988

By S. J. Jung

The stability of underground openings for deep mines is critical, since incorrect orientation or geometry of shaft, stope, and accessways often results in the failure of openings due to high stress concentrations. Current state-of-the-art procedures for the design of underground openings are largely based on rules of thumb, experience, and trial and error. Rules of thumb are invariably based on the past experience of the designer and may not remain optimal as the depth and complexity of mine workings increase. Often, as conditions change, and as the interaction between geology, in-situ stress conditions, and mine openings grows more complex, the cost of maintaining structural integrity becomes prohibitive. Existing design procedures for deep mines are rapidly becoming inadequate as mines are worked at greater depths. For example, in the Coeur d' Alene Mining District, several million dollars are spent for inspection, repair, and maintenance on a regular schedule, largely as a result of inadequate design. In these types of situations, it is of critical importance that a new design methodology be developed for the layout and design of underground openings.

Jan 1, 1994

By Keith Kern, Joseph P. Matoney, Kenneth E. Harrison

INTRODUCTION "When there is an economic need for cleaning fine coal to low ash, the technical difficulties will be solved. The utilization of fines below 28 mesh will determine the degree to which beneficiation will be carried. Whenever these "fines" will be processed into smokeless fuels or used as new materials for the chemical industry or for hydrogenation, a higher degree of cleaning will be economically justified”.(l) This quote from the 1943 edition of Coal Preparation correctly anticipates the use of fines as "smokeless" fuel or as a future component of boiler fuel and a1 so as a raw material for the chemical industry or possible new uses such as for internal combustion engines. Fines utilization depends upon an economic need and possibly neglects to say an environmental need or a government funded development program as the prime mover for development. In any event, this conference is faced with the "need" to beneficiate fine coal and undoubtedly among the processes that will be discussed are the unit processes of future coal preparation facilities. HISTORY OF DEVELOPMENT It is interesting to look at where we have been, where we are now and where we possibly will be in the future. In order to put some limitations on this presentation, we will for the most part confine our remarks to this century and particularly developments since World War II. We also will limit our remarks to the preparation of bituminous coals in the United States, and for the most part, to wet processes. In the 1943 edition of Coal Preparation, clues to the processes being used to beneficiate coal are contained in the Chapter Titles. An extensive treatment of "Hand Picking" was contained in Chapter 8. Chapter 7 contained a description of methods for "Preparation at the Face". Chapter 18 dealt with "Slurries, Sludges, Slimes and Water Treatment" and defined this size of coal as being the size at which "Jigs and Chance Cones are usually dispensed with and other methods such as the hydrator, tables, and froth flotation are employed. Generally the limit is 1.7mm to 3.4mm. In the days before World War I1 the then modern bituminous coal preparation plant contained jigs or Chance Cones for the coarse material +6.35mm, concentrating tables, launders or some classifier type of coal washer for the fines (6.35 x 0.15mm), and froth flotation as the only method of cleaning minus 0.15mm coal which was used very sparingly.

Jan 1, 1988

By Cheryl Lee Vatter

The Lupin mine consists of a gold mining, milling, and refining complex located 400 km (250 miles) northeast of Yellowknife and 100 km (60 miles) south of the Arctic Circle on the southwest shore of Contwoyto Lake in Canada's Northwest Territories. (Fig.1). The mine was commissioned in 1982 and is presently producing 6 t/a (193,000 oz per year) of gold from 612 kt (675,000 st) of ore. The success of this operation is due to such factors as the continuity and grade of the ore body, the competency of the host rock, low mining costs, efficient milling, the transportation of people and materials to and from the mine site, and the unique work schedule resulting in a stable workforce. Echo Bay Mines operated a silver mine at Port Radium on Great Bear Lake, Northwest Territories, from 1965 to 1982. The company first obtained an option on the Lupin property from Inco in February 1979 and completed an underground exploration program in 1979-1980. Topographic relief is low and vegetation is sparse in the continental subarctic climatic zone, consisting mainly of moss and lichens. Temperature extremes are from 24°C to - 45°C (+75°F to -49°F) with an annual mean of -12°C (+54°F). Permafrost extends from near surface to 500 m (1640 ft) in the ore zone. Its remote location and harsh climate presented some challenging design and logistical problems. Construction began at the Lupin mine site in 1980. Before construction, a 1.5-km (5000-ft) gravel landing strip was prepared, suitable for landing a C 130 Hercules. The entire Lupin project took 20 months to build. All of the men and materials were transported to the site with some 1100 Hercules flights and several hundred Convair 640 flights. The Convairs transported construction crews, which numbered 400 at their peak, and also carried 3.2 kt (3535 st) of supplies during construction. The facilities were constructed and commissioned for a total cost of C$135 million. The operation was originally designed to throughput 860 t/d (950 stpd). Expansion in 1983, circuit refinements, and some capital projects brought the daily throughput to 1.7 kt/d (1850 stpd). The underground mine delivers 612 kt/a (675,000 stpy) of ore to the mill at an average head grade of 8.46 g/t (0.3 oz per st). The ore is nonrefractory and is processed in a conventional cyanide leach using the Merrill Crowe process. Gold recovery is about 95.0%. The average production cost is $US5.85/g ($US 182 per oz) based on 1987 figures. Mining at Lupin – Geology The Lupin deposit occurs in amphibolite grade iron formation overlain by mudstones (phyllites) and underlain by graywacke (quartzites). Contacts between the wallrock units and the iron formation are well defined. It has been folded and tilted into a megascopic antiform-synform-antiform structure (Fig. 2). Gold occurs primarily within the sulfide rich iron formation, with some minor occurrences in sulfide poor iron formation. The distinction between sulfide rich and sulfide poor iron formation is based on a visual cutoff of 5% total sulfide content. Mining widths are determined by an assay cutoff of 4.2 g/t (0.15 oz per st) gold. There are few tons between 1.7 and 4.2 g/t (0.06 and 0.15 oz per st). The amphibolitic iron formation at the mine ranges from 1.5 to 20 m (5 to 65 ft) wide and has been followed over a strike length exceeding 1.7 km (5600 ft). The wider portions of the ore body tend to occur at its north extent and south nose. The gold-bearing iron formation appears on plan as a Z-shaped structure made up of three zones: the West, Center, and East. The West and Center zones dip steeply to the East (75° to 90°). Each of the ore zones plunge at an angle of about 65°. Total strike length of the three zones is more than 610 m (2000 ft). The zones are confirmed at a depth of 650 m (2130 ft) below surface. The Center zone is the widest and varies from 4.5 to 20 m (15 to 65 ft) while the West is the narrowest, averaging 1.5 m ( 5 ft). The footwall is comprised of quartzites that are strongly jointed and locally grades into phyllite, which comprises the hanging wall. The hanging wall and footwall are reversed in the West zone. Mineralogy of the ore at Lupin consists of amphibole minerals (hornblende, cummingtonite, and grunerite), feldspars, quartz, occasionally garnet, pyrrhotite, arsenopyrite, minor pyrite, and trace chalcopyrite. Also found in minor amounts are scheelite, apatite, epidote, calcite, tourmaline, and some arsenides (notable loellingite). Quartz

Jan 1, 1989

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

Jan 1, 1987

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

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

Jan 1, 2002

By Hasan Sevim

A computerized heuristic model for an efficient and economic design of belt haulage networks in underground coal mines is discussed. The model is versatile, easy to use, and is valid for all mining methods singly or in combination. It may be used to design new networks or to analyze existing networks. The input data required are 1. maximum production rates from a section, 2. probabilities of production rates, 3. equipment costs, and 4. data necessary for network coding to represent different cases. The objective function is the "expected minimum cost per ton of coal hauled" while meeting production requirements. The output includes belt width, motor horsepower, bin size etc. and also values of operating parameters for each belt such as belt speed, effective tension, power consumption etc. A detailed account of the cost involved is also provided for each unit as well as for the entire network. Description of the model and an example of the design of a belt network in a mine with two longwall and four development sections are presented.

Jan 1, 1986

Underground mining has the potential to be one of the more dangerous industries in the United States and the world. Following recent disasters in Sago, WV; Darby, KY, and Crandall Canyon, UT, the mining industry was reminded of the need for a system to allow trapped miners awaiting rescue to survive following a disaster. To address this need, Battelle has developed a Mine Barrier Survival System (MBSS) that surpasses all current technologies and sets a new standard for keeping miners alive in the event of a fire, flood, explosion or collapse. The MBSS creates a temporary refuge for miners who cannot escape by using inflatable barriers and a post-fire atmosphere scrubbing system to define the ?safe haven? and remove CO2 and CO while replacing oxygen and providing heat. Current personal-protective systems in use are designed to provide breathing air for only a few hours when, realistically, it may take days for rescue teams to reach trapped miners. Battelle?s MBSS is designed to aid up to 15 miners for four days.

Jan 1, 2008

By M. D. Banghart

The Republic of South Africa is an outstanding example of the benefits to a new land by the application of individual enterprise, capital and know-how. While South Africa was colonized by European and Bantu people some 300 years ago, it was only 70 to 80 years agog when diamonds were discovered at Kimberley and gold on the Witswatersrand, that the mining, industrial and commercial potential of the country began to emerge. South Africa s national income rose from R1, 200 million in 1945 to over R4 billion in 1960 (1R = 10 shillings = $1.40 U. S.) The annual national income per head of population -- including all races -- is currently R254 (£127) This compares with £24 for India., £30 for Nigeria, and £50 for Ghana. Allowing for the increase in population, the increase in real income per head of population was about 40% during the past 15 years.

Jan 1, 1962

By D. Yeager

Researchers at the U.S. Bureau of Mines have successfully used a genetic algorithm to tune computer models of three column flotation units simulating a separation circuit. The computer models rely on empirical constants that are selected so that the models accurately predict the performance characteristics of individual column flotation units. Unfortunately, these empirical constants have traditionally been selected using a time consuming trial and error process. A new and more efficient approach to selecting empirical constants for computer models has been developed. A genetic algorithm has been used to select the empirical constants for column flotation models used to simulate a three column circuit. Genetic algorithms are search algorithms based on the mechanics of natural genetics. They rapidly locate near optimum solutions in difficult search spaces. Bureau researchers have used these innovative search algorithms to dramatically reduce the time needed to select empirical constants for column flotation models. The approach is applicable to the selection of empirical constants for numerous other computer models.

Jan 1, 1995

By Karl Ingmarsson

Larger mines spend millions of dollars annually on rock drilling tools, but often purchase based on historic consumption. The results vary from excess inventory to acute tools shortages. These lead to emergency sourcing, sometimes mixing products that don’t perform optimally. Boart Longyear analyzed the usage pattern of 96 thousand drill bits at thirteen different underground mining contracts, showing very high variability. The monthly standard deviation on a product level was 42%, with the highest for development and low volume bits. A new rule of thumb is suggested to carry only one month of safety stock for high volume bits, but increase to three months for low volume bits. This balance will reduce the risk of outages with less inventory. The accuracy can be improved by considering the specific conditions at each mine and reclassifying individual bits for a different safety stock level. Root cause analysis and collaborative forecasting between mines and suppliers can lead to further optimization of availability and inventory levels.

Feb 1, 2024

By D. Loy LeRoy

As we are all aware, the impacts of the 1977 Surface Mining Act and the implementing OSM Regulations require application of new, as well as existing, equipment and technologies. Our firm recently completed (June, 1979) a study for the Department of Energy titled "Analysis of the Impact of Public Law 95-87 on Mining Performance." The objective of the study was to analyze the impact of the Act and its regulations on mining performance in order to provide recommendations to alleviate the impacts identified, including identification of research needs. This paper presents some of the highlights of that report: projected cost increases; additional equipment needed for compliance; review of methods currently practiced or those which can be modified; and will focus particulary on equipment and technology applications.

Jan 1, 1980

By W. G. Kegel

The past few years have produced a rash of demands for changes in America's way of doing things. From manufacturer to consumer, the rules have been changed. Some of the demands were needed and, when legislated to requirements, did manage some good for all concerned. Other demands made by zealots and taken up by regulatory agencies have forced costly changes on our way of life. No segment of industry has been exposed to more new rules and regulations than the deep coal mining industry in the United States. There is no denying that some good has evolved, but most mining men know that the effects have done more to reduce production and increase cost than they were ever intended to do.

Jan 1, 1975

By D. W. Fuerstenau

One of the more energy-efficient methods for size reduction, namely particle-bed comminution under high compressive loads, was employed to grind eight minerals and ores, using up to 250 MPa pressure. A new approach based on the self-preserving character of the ground product size distributions and a linear energy-reduction ratio relationship is proposed for analyzing this mode of size reduction. The grind ability of the material when comminuted in a particle bed is about one-half that of the single-particle crushing mode, which is in agreement with previous estimates of the relative efficiencies of energy utilization based on direct measurements of surface area produced. The loss of efficiency may be attributed to energy dissipation by particle interaction phenomena in compressed beds.

Jan 1, 1993

By George C. Carter

It was just a little over three years ago that the "Go" button was pushed to start operations on the blending piles at Oro Grande. As is usual with any new installation, there were problems that had to be solved. This paper will concern itself with those problems directly related to quality control. To obtain a consistent outgoing quality level at the most economical level for a finished product requires two main points: 1. Measurement of variations in the process, 2. A method of making a predictable change to adjust for variations. The manufacture of cement requires a relatively high degree of accuracy if a consistent outgoing quality level is to be maintained. The blending piles were our solution to this problem.

Jan 1, 1962

By F. W. DeVries, J. Brent Hiskey

The search for noncyanide lixiviants for treating gold and silver ores and concentrates is not a new endeavor. However, in recent years, the overall activity to seek a less environmentally controversial reagent than cyanide has occasioned considerable developmental work. While the elegant economy and simplicity of cyanide, because of its selectivity and the ease of recovering gold and silver from the cyanide complexes, is well recognized, two failings have justified the search for alternates: increasing concerns and costs to assure avoidance of environmental damage from cyaniferous liquids and solids, and the need to simplify recovery of gold and silver from refractory ores. Generally, process development work to date on the alternate reagents has lacl<::ed environmental focus ( Hiskey and Devries, 1992). This paper covers briefly the chemistry of various alternates to cyanide and reviews their environmental implications. In each case, environmental overtones are reviewed following the chemistry. For ease of follow-up, the authors have listed sources of information in detail. Alternates examined include:

Jan 1, 1992

By Randy Fulton, Ignacy Puszkiewicz, George H. K. Schenck

Stuart Olson Construction, Inc. excavated 385m (1263 ft.) long twin tunnels with 6.25m (20.5 ft.) O.D., a portal and a cross-over for the Edmonton, "Metro". The work for Contract No. 76458, SLRT South Tunnel and Portal, in Edmonton, Alberta, Canada was performed in 1989 and 1990. NATM was specified because of the instability of glacial deposits. The contractor decided to use "two firsts" on this project: Fully mechanized NATM and a novel excavator¬mounted tunneling machine. An excavator was equipped with a cutter boom, and a loading and conveying system. This novel machine cost only about one-third of the price of a special NATM roadheader. This new combination of NATM and an excavator-mounted cutter boom offered a cost-effective alternative to shields, slurry TBMs and roadheaders for short tunnels in unstable ground. For NATM tunneling in hard rock or in mixed faces, the roadheader-type cutter boom can be replaced with an impact hammer (hydraulic breaker).

Jan 1, 1993