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By Karel SPIRKO

The one-dimensional mathematical model of unsteady-state heat conduct ion through multilayer (wall and floor of airway consist of several layers of material and rock) and two-surface (dry and wet surfaces) airway wall and airway floor with convective heat and moisture transfer between their surfaces and mine air with time dependent changes of temperature and humidity has been developed on the basis of introduction of acceptable simplifications of description of heat and moisture transfer between rock mass and mine air. The analytical solutions for calculation of the surface temperature of any type of surface of airway wall/floor and for calculation of water vapour pressure near the wet surfaces at given time are obtained by means of the model presented. These solutions are illustrated for the case of general harmonic course of air temperature and humidity. Balance of heat and moisture transfer between surfaces of airway wall and floor and mine air is solved by an iterative method in the partial sections of airway. The temperature and the humidity of mine air on the outlet of each partial section of airway at given time are its resolutions and they are the input data for following partial section. By this way it is possible to solve complicated dynamic changes of mine air temperature and humidity in an air way system by relatively simple mathematical apparat.

Jan 1, 1991

By J. H. Tu

The technical paper correctly points out that the kriging variance is not a good measure of the uncertainty of the estimated (i.e., kriged) value of individual blocks. The authors claim that their proposed jack-knife method, which is a rekriging of each block by eliminating, in turn, one sample from m the original sample set and then taking the average of the rekriged estimates, not only gives good block estimates, but the resulting jackknife kriging standard deviation is a useful indicator of the "true uncertainty associated with block estimates." However, they immediately abandon the idea of using the block-by-block standard deviations, reasoning that these standard deviations are not independent and that there is no easy way to utilize them. There may be another reason for not using them. The jackknife standard deviations for individual blocks given in their example are mostly in the range of 0.004 to 0.005 oz/st (0.14 to 0.17 g/t) with only one block having a high value of 0.012 oz/st (0.41 g/ t). These individual block standard deviations are as low as the jackknife standard deviation for the mean grade of the entire shape, i.e., 0.0041 oz/st (0.14 g/t). Do they represent the "true uncertainty" .of the individual block estimates? Could the authors explain this? In a global shape consisting of a large number of blocks, any given sample will affect the kriged estimate of only those few blocks within its vicinity. This is the rationale for the authors' selective rekriging, making the jackknife algorithm more efficient. On second thought, why not do away with jackknifing altogether? Just cumulate and normalize, if necessary, the kriging weights of each sample used during the ordinary block kriging process, and then compute the global variance from these kriging weights and their respective sample grades? After all, isn't the global mean grade nothing but the weighted average of the samples used in the estimation? Reply by G.S. Adisoma and M.G. Hester The jackknife is one of the many tools in a practitioner's toolbox to solve estimation problems. The strengths of the technique lies in its simplicity, i.e., it uses the concept of mean and standard deviation and the fact that it can be easily combined with other tools, in this case kriging. Because the jackknife kriging (JK) estimate is also the mean of the pseudovalues, the JK standard deviation is attractive just as the standard deviation of the mean explains the variability of the data. The difference is that the pseudovalue calculation in jackknife kriging uses the ordinary kriging (OK) weighting scheme instead of simple arithmetic averaging. The data used to illustrate the jackknife technique in the paper con¬sist of high values that are roughly three times the low values. The resulting JK estimate of the block grades show that the highest estimate is roughly twice the grade of the lowest estimate. The contrast between the low and the high estimate is more evident in the JK estimate than in the OK estimate, even though the mean grades of the blocks for the two estimates are very similar. Nonetheless, in this paper, we are concentrating more on the need for a more realistic measure of uncertainty, or precision, for the estimate. Unlike its OK counterpart, the JK standard deviation of the blocks clearly reflects the original data variation. The highest JK standard deviation of the blocks is three times its lowest value. This follows our intuition that, when the samples used to estimate a block is more variable, the resulting estimation variance (or standard deviation) should be higher than the case where the samples are more uniformly valued. However, block-by-block standard deviation or variance is of little practical value in reserve estimation and classification, as well as in mine planning. One is usually more interested in quantifying not the variance of the individual block estimate, but the uncertainties associated with a much larger dimension, such as the minable reserve. Thus, the thrust of the paper is to find a simple way to obtain a single estimation variance or standard deviation associated with the reserve grade estimate. The discussion by J.H. Tu did not mention how one would obtain the global variance from the OK weights and the sample grades. As a technique that offers a data value-based measure of uncertainty for its estimate, the "leave-one-out" jackknife fills this need nicely through the block kriging shortcut approach described in the paper. Note: The first column and the last two columns of Table 3 in the paper should have contained a single number each, namely, an OK estimate of 0.0317, a JK estimate of 0.0333 and a JK standard deviation of 0.0041 oz/st, respectively, for the shape, as are obvious from the text.

Jan 1, 1997

By C. A. Fleming, C. J. Ferron, E. S. Makande, P. Mason

Several small mines are producing refractory gold concentrates in Zimbabwe and, owing to the small scale of most of these operations, their output is processed in a central custom roaster in Kwekwe. Owing to the wide variations in the feed to the roaster, the efficiency of the old Edwards roaster is deemed unsatisfactory with gold recovery from concentrate as low as 70%, and alternative processes were evaluated to improve this situation. Various processes for oxidizing refractory arsenopyrite/pyrite concentrates to liberate gold prior to cyanidation are known. In practice, only two have to date met with wide industrial acceptance, pressure oxidation and roasting. Owing to the ·high arsenic and antimony levels of some concentrates in Zimbabwe, the only viable roasting alternative would be two-stage air-roasting of concentrates in two reactors, as practiced at Giant Yellowknife, Campbell Red Lake and New Consort. This process, however, creates an arsenic disposal problem, and produces gas of fairly low SO2 content, requiring a large gas handling system for acid production. On the other hand, there are currently four commercial pressure oxidation plants treating refractory gold ores (McLaughlin, Mercur, Getchell and Goldstrike) and three treating concentrates (Sao Bento, Porgera, Campbell). One other concentrate oxidation plant is being constructed (Con), and several more are being considered.

Jan 1, 1992

By L. Yuan, J. H. Rowland, R. A. Thomas, D. Bahrami

Blasting is the main method of production in many non-coal underground mining operations and produces multiple toxic gases as a result. The Mine Safety and Health Administration (MSHA) requires mine operators to measure the level of toxic gases in mines as frequently as necessary to ensure they are below regulatory safety limits. The current practice uses portable gas monitors to check the concentrations of toxic gases after a fixed post-blast time. This paper studies the application of a gas monitoring system in the return entry of a limestone mine to determine a safe re-entry time. The National Institute for Occupational Safety and Health (NIOSH) conducted such a monitoring program in a limestone mine from September 2016 through May 2018. NIOSH/PMRD (Pittsburgh Mining Research Division) is endeavoring to develop workplace solutions to improve detection of and reduce the risk of hazardous conditions. This study showed that the use of gas monitoring in the return air can be a useful tool at the mine operator’s disposal to detect and reduce the risk of hazardous conditions and also to reliably estimate the re-entry time. INTRODUCTION Underground miners are at risk of overexposure to toxic gases from various sources. Production blasting operations using explosives in underground mines are a source of multiple toxic gases that could be harmful and even fatal to humans if prolonged exposure occurs beyond the safe limits. According to Code of Federal Regulations, Title 30, Part 57.5002 [1], metal and nonmetal mine operators must conduct gas, mist, and fume measurements as frequently as necessary to determine the adequacy of control measures to protect miners from overexposure to harmful gases. MSHA will inspect mine operators’ activities to verify the evidence of such measurements [2,3]. Continuous gas monitoring in the return air is an effective way to make sure that toxic gas concentrations at the working areas are kept below the regulatory limits before the return of miners. Toxic gases can enter the mine atmosphere from different operations, such as welding, combustion, and blasting. The main gases produced from blasting are carbon monoxide (CO), carbon dioxide (CO2), nitric oxide (NO), and nitrogen dioxide (NO2). Other gases are produced in small amounts and, therefore, were not considered for this study. Related regulatory and recommended exposure limits of CO, NO, and NO2 gases are summarized in Table 1 [4].

Jan 1, 2019

By Gu Guohua, Hu Yuehua, Zhang Qin

The flotation behavior of pyrrhotite was investigated by using ethyl xanthate as a collector. Results showed that pyrrhotite had good floatability from pH 2 to pH 11, and poor floatability when pH was >12. The flotation of pyrrhotite was dependent on pulp potential on various pH values. The potential-pH range for pyrrhotite flotation was established. Cyclic voltammeny and FTIR spectroscopic analysis showed that the major adsorption product of ethyl xanthate on pyrrhotite was dixanthogen. The intensity of FTIR signals of dixanthogen adsorption on pyrrhotite and the anode current of a pyrrhotite electrode and flotation response of pyrrhotite are correlated with pulp potentials.

Jan 1, 2005

By Jean E. Kiel, Rob Dorey, Dirk J. A. van Zyl

1.1 INTRODUCTION This chapter seeks to provide an overview of important factors involved in the implementation of heap leach technology as a vehicle for gold extraction from its ores. Toward this end, brief discussions of the various elements important to this process are presented here. Topics covered in this chapter include a brief history of the process; basic chemical reactions responsible for the process ; characteristics of ores amenable to heap leaching; location of the greatest concentration of heap leach projects in the U.S.; heap leach components, methods, regulatory/permitting considerations; and economic considerations. Most of these subjects are treated at much greater length in subsequent chapters. Thus, this chapter serves as a summary for those which follow, in addition to providing the reader with a feeling for those topics which may be of most immediate interest. 1.1.1 History of the Process Modern Day Precious metal heap leaching technology has developed over the past decade, although the principles of heap leaching per se, as well as those of gold extraction through cyanidation have a long history (see discussion below). However, it is in the past 10 years that heap leaching has developed into an efficient method of treating oxidized gold and silver ores. It has proven to be both an efficient way to extract precious metals from small, shallow deposits, as well as an attractive way to treat large, low- grade, disseminated deposits. Heap leaching has several advantages compared to conventional milling (i .e., crushing, grinding, and agitation leaching). In general, these advantages include simplicity, lower capital and operating costs, shorter start-up times, and less intensive environmental regulatory concerns. A possible disadvantage of currently existing heap leach technology is a potentially lower percentage of metal extraction being obtained from the ore than would be the case with conventional milling (Potter, 1981; Hiskey, 1985; and Thorstad, 1987) The actual principle of heap leaching has a long history as inferred previously. For example, mines in Hungary recycled copper-bearing solutions through waste heaps in the mid-sixteenth century (Hiskey, 1985), and Spanish miners percolated acid solutions through large heaps of oxide copper ore on the banks of the Rio Tinto around 1752. By 1900, leaching operations were employing such techniques as leach/rest cycles to maximize copper recovery (Taylor and Whelan, 1942; Thorstad, 1987). Copper dump leaching is currently

Jan 1, 1988

By D. -P. Georgescu

A native sulfur mineralization associated with pyrite and iron oxides containing 20% iron was quarried in Calimani Mountains, Oriental Carpathians. The wastes resulted from this exploitation activity, with a volume of 33 millions m3, were stored in four dumps while the wastes resulted from ore processing using flotation and the subsequent treatment were disposed of in a settlement pond. The investigations carried out in the entire mining area focused on the following two aspects: - the impact of environmental factors mainly on the hydrographic network; - the characterization all pollution sources, namely dumps, leachates and the quarry and the assessment of acidic drainage generation. The total sulfur content of the excavated ore was 3.8% while in dumps it ranged between 0.54 and 1.06%. Analysis of samples collected along the hydrographic network, draining the mining area, resulted in pH values ranging between 1.7 and 2.83 as well as high concentration of iron (4525 mg/L to 1.6 mg/L), copper (1.87 mg/L to 0.04 mg/L) and zinc (13.13 mg/L to 0.3 mg/L). Samples from the quarry, the dumps and the tailings pond were analyzed in order to deter-mine: - the maximum potential acidity (MPA); - the acid neutralization capacity (ANC); - the net acid production potential (NAPP).

Jan 1, 2005

By V. B. Achanti, A. W. Khair

This paper deals with the effects of various bit geometric parameters on the failure characteristics of rock material. Improving the rock-cutting efficiency of continuous mining machines has proven to be an interesting research topic during the last few decades. In the current study, an automated rotary rock-cutting simulator, duplicating a continuous miner, was used to study these parameters in the laboratory. The improvement of bit performance in terms of its life, cutting efficiency, dust production and energy consumption has drawn widespread attention. Research has demonstrated that optimizing bit parameters can result in efficient fragmentation of rock. Indiana limestone rock material was used in this study. The bit tip angle, bit tip size, bit spacing, depth of cut and cutting velocity of the rotary drum were varied to investigate the failure characteristics. The cutting pressure, penetrating pressure and the depth of cut were monitored on the simulator during the cutting process. Airborne dust and settled dust were measured for each experiment. Then, the specific airborne dust and specific energy were determined for each type of bit. Triaxial tests were carried out to correlate the failure angle of the rock in the static case to the dynamic cutting process failure. The physical and mechanical properties of the rock were also determined to assess their influence on the wear characteristics of each bit.

Jan 1, 2000

By R. A. Hodge, J. W. Murray

By an order of the Executive Council of the Provincial Government dated January 18, 1979, a Royal Commission was established under the Public Inquiries Act to: 1) Examine the adequacy of existing Federal and Provincial requirements for the protection of the health and safety of workers associated with exploration, mining and milling of uranium in B.C. and for the protection of the environment and of the public. 2) Receive public input on these matters. 3) Make recommendations for setting and maintaining standards for worker and public safety and for the protection of the environment as a result of the exploration for the mining and milling of uranium ores. At that time,and at the present time in British Columbia, no uranium mines were, or are, in operation. Exploration at a number of sites was under way and one deposit, the Blizzard deposit, with estimated reserves of 1.92 million tonnes containing approximately 4.74 million kilograms of U308 was at an advanced stage of exploration. With our Terms of Reference, and because of the diverse geology of the western Cordillera which could potentially host a broad variety of uranium, deposit types, the Commission was faced with the task of considering environmental and health protection necessary for the potential development of [any] kind of uranium mineralogy situated in [any] of the environments found in British Columbia. Over the past 17 months since that time, much has transpired, culminating in the February 27, 1980 Provincial Government announcement of a seven year moratorium on exploration, mining and processing of uranium coupled with the termination of our Inquiry before completion of the hearings. The Government sebsequently modified its decision to terminate the Commission completely, deciding instead to stop the hearing but allow time for the Commission to gather in the work that had been initiated

Jan 1, 1980

By A. E. Prokop, J. P. Rider

Previous research on continuous mining operations has shown that significant interactions exist between dust-control parameter application and the resulting respirable dust levels. However, simply increasing the level of the controls does not guarantee a reduction in respirable dust levels. Full-scale laboratory tests are being conducted to provide information to longwall operators that would assist them in selecting control parameters to reduce dust levels for mine-specific conditions. The interactions between face air velocity, shearer water quantity, drum water spray pressure, external water spray pressure and spray system design were evaluated in a simulated 2.13-m (7 ft) coal seam for two cutting directions. Locations around and downwind of the shearer were monitored to evaluate relative changes in respirable dust levels as a function of each control parameter.

Jan 1, 2003

By R. A. Jankowski, J. J. McClelland

Laboratory tests were conducted by the U.S. Bureau of Mines to determine the respirable dust reduction effectiveness of and interaction between face airflow and water sprays for a continuous miner. Exhausting face ventilation was varied from 3,000 to 9,000 cfm (1.42 to 4.25 m3/sec). Spray-water-flow was varied from 15 to 35 gpm (0.057 to 0.132 m3/min), while nozzle operating pressure was varied from 80 to 200 psi (5.62 to 14.06 kg/cm2). Results from gravimetric dust samples indicated that airflow had the greatest individual impact on reducing dust levels, with return and operator concentrations reduced by as much as 57% and 99%, respectively. Interactions between dust control parameters were found to be significant. Often, these interactions defined a level of application for a control parameter where further increases in that parameter failed to produce additional reductions in dust concentrations. In several cases, increases in a control parameter resulted in higher dust levels at the operator's location. These higher dust levels were attributed to additional rollback and/or airflow turbulence. Regression modeling indicated that increases in airflow to 8,400 cfm (3.96 m3/sec), water flow to 25 gpm (0.095 m3/min), and water pressure to 140 psi (9.84 kg/cm2) typically were beneficial at the operator and return sampling locations.

Jan 1, 1991

By Xinghai Lei, Nan Wang, Xiangyang Jin, Yuqi Ren, Yang Li

The upward mining method, as an effective mining method, is widely employed in close-multiple coal seams to extend the service life of coal mines and improve the recovery rate of residual coal resources. However, the upward mining feasibility of the upper coal seam is related to the height of caved zone of interlayer strata after lower coal seam extraction. Therefore, it is extremely important to determine the height of the caved zone after the lower coal seam. The No. 9 coal seam (upper coal seam) and the No. 12- 1 coal seam (lower coal seam) in the Qianjiaying coal mine are selected as the research object. The characteristics of the interlayer strata structure field, fracture field, and subsidence field are studied using UDEC numerical modeling, in which the height of the caved zone is gotten after the No. 12- 1 coal seam extraction. Furthermore, the height of the caved zone is verified by onsite detection of GPR and BTVI. In the numerical modeling, the height of the caved zone is 19.6 m. And the height of caved zone is 14.7 m and 15.4–21.2 m, respectively, by using the onsite detection of GPR and BTVI. The heights of the caved zone onsite detection results are consistent with the numerical modeling result. The No. 9 coal seam is always located above the caved zone after the No. 12- 1 coal seam extraction. Consequently, the upward mining of the No. 9 coal seam is feasible.

Jun 28, 2023

"?irovski Vrh" Uranium deposit in Slovenia is type of mineralized sediments in ore bearing belts of deflected S-structure, in which the belt is less than 25% mineralized. Ore bodies of irregular small thickness, unreliably contoured, made difficult estimation of correct contours, which are necessary for design and exploitation works. The deposit was explored by vertical boreholes made at the surface and by boreholes fanned from drives. For the purpose of application of this estimation method, 1 m composites have been made in slices which were parallel to the referential belt plane. Estimation was made for 1,5m x 1,5m x lm blocks which do correspond to the size of the containers in the ore transport stage. Two variables have been used: the first, indicator of function (ore=1, waste material =0) and kriging method; the second, logarithmical content of Uranium and Minimum Mean Square Error Estimation (MMSEE) method. Based on estimated values of those variables, probability of ore quantity as well as Uranium content over determined cut-off were calculated for each block. Those two parameters enable determination of the most probable form of the ore body in ecah one metter slice. Estimation of sixteen small blocks in a large 1,5m x 1,5m x 1m block made possible estimation of ore quantity and its content for the determined cut-off, as well as categorization of exploration. This method is being used for estimation of ore reserves and planning of exploration operations. The described' methodology was tested in an exploitation block.

Jan 1, 1986

By Ora H. Rostad

Some gold production was obtained from placer deposits in Puerto Rico during the period of Spanish colonial rule, but most of the gold was produced prior to 1536. There had been no significant hard rock mining, other than in quarries for road metal or construction purposes or for raw materials to make cement, prior to 1954 when the government of the Commonwealth of Puerto Rico recognized that changes in the mining code were necessary to stimulate exploration for metalliferous ore deposits. The Puerto Rican Legislature passed an act (No. 6), which was approved on Oct. 6, 1954, that reiterated that the title to all "commercial minerals" is vested in the Commonwealth of Puerto Rico. The act provided for the establishment of a Mining Commission, stipulated the power% with which the Mining Commission was vested, and provided that no one may dispose of "commercial minerals" in Puerto Rico without a Commonwealth grant issued on its behalf by the Mining Commission. Act No. 6 also provided for prospecting permits, mining leases, and grants. The Mining Commission, by authority granted to it by Act No. 6, adopted regulations effective Nov. 21, 1957, relating to prospecting, leasing, and producing commercial minerals in Puerto Rico. Royalty rates were also specified for various types of deposits at different annual tonnage production levels.

Jan 1, 1991

By D. A. Knight, V. G. Medina

Kidston's 14 000 t/d mine and mill complex in Queensland, Australia, started operation in January 1985. During its first 18 months of operation, the mine produced over 7000 kg (225,000 oz) of gold, making it the largest gold producer in Australia. This paper describes the planning, logistics, flowsheet, and process for the Kidston project, and some of the problems encountered during startup and the measures taken to correct these problems.

Jan 1, 1987

By Paul M. Godlewski, Gerry Millar, Harvey W. Parker

Recently, over 70 railroad tunnels in the western USA have been enlarged and rehabilitated "live track," utilizing modern excavation and lining methods, in most cases to clear the larger double-stack container cars. This paper presents the general conditions of the tunnels and the technical aspects and costs of the rehabilitation of approximately 40 kilometers of tunnels, which range in length from a few hundred meters to 13 kilometers. Concrete-lined tunnels were rehabilitated by grouting and rock bolting, while unlined or timber-supported tunnels were relined with shotcrete and rock bolts. High output lining methods were required because of limited track time. Enlargement has been accomplished primarily by carving notches out of the concrete linings with a roadheader, or occasionally with saws. Unlined tunnels were enlarged by drill-and-blast or roadheader excavation.

Jan 1, 1991

By S. K. Kawatra, R. A. Seitz

Nonpolar hydrocarbon oils are commonly used in the flotation of minerals to fulfill a number of roles. These roles include reducing overall reagent costs by replacing part of the collector addition, improving coarse particle recovery without reducing selectivity, as well as several others. This paper will cover the types of oils which are available, their physical and chemical characteristics, and the effect of these characteristics on the function of hydrocarbon oils in flotation. In addition, techniques for improving the behavior of oils will be discussed.

Jan 1, 1986

By James A. Pappas, Francis L. Fong, Tracy K. Lundin

The impact of environmental regulations on the quality of tunnel groundwater discharges is examined in this paper. Tunnel contractors usually depend on specialty contractors or outside consultants to deal with this issue without carefully examining their activities to minimize the production of contaminants. This approach may not be adequate given increasingly more restrictive environmental requirements. This paper surveys the general practice of the tunneling industry and attempts to identify sources of contaminants which may affect the total suspended solids, oil and grease content, and pH of the discharges. Strategies for treating the discharges are included. Field data and observations from several tunneling projects are also presented.

Jan 1, 1995

By R. D. Begley, A. W. Khair

This paper presents an assessment of modeling techniques in predicting subsidence over longwall areas. Extensive instrumentation has been made, both at the surface and subsurface, at three longwall mines in northern West Virginia. Ground movements were monitored during a three- year period until subsidence ceased in these areas. A comparative analysis of empirical subsidence prediction techniques including graphical (NCB), profile function, influence functions (USBM, Bals', Knothe's, and zone area) will he presented. A brief geologic and topographic description of each site, along with the instrumentation and monitoring program, will also he presented.

Jan 1, 1990

By M. Gershon

The article is an excellent one, providing a basis for the conduct of the major tasks in underground mine planning and scheduling. I object to only one word. Since that word is featured in the title, it is a strong objection. The word used erroneously is optimization. There is no optimization performed in the article. By definition, an optimization is an approach that yields the best of all possible solutions. We can limit this definition to its narrowest sense, the optimization techniques of operations research such as linear programming, or broaden it to include less sophisticated search procedures that find the best mine plan or schedule. No such claim is made in the article. No optimization of any kind is performed. The major tool used is a simulation, which, in truth, is probably more appropriate for this problem than any optimization. But simulation does not optimize. It only allows for quick evaluation of many options, leading to the selection of the best plan "of those tried," not "of all possible plans." This article, of course, is not the first mining paper to make this error. It seems that the word optimization is used more to impress people than to refer to any act of optimizing. It is used in a way that implies "try to optimize" or "come close to optimize," but not to really do it. It is usually appropriate to quietly ignore this error. But when Dr. Chatterjee, who has done so much good work in true optimization projects and whose work I have come to admire, makes such an error, it is time to publicly dispute it. Having discussed this very point with him in the past, I am sure that he will agree. If he can support this position in his response, perhaps people will start to listen and we can reduce the number of false claims of optimization. This should be easy for him in this case since the approach he used, simulation, is the more appropriate one anyway.

Jan 1, 1987