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By Wolfgang Güter, Paolo Romualdi

Current extensions of the European railway network require long TBM drives in the diameter range of 10 m under critical and varying rock conditions. In view of that and in particular for a 7 km long high-speed railway tunnel in Spain to be started in mid-2003 a new design for a hard rock Double Shield TBM has been developed by SELI (Italy) and Jäger (Austria) in collaboration with Mitsubishi (Japan)and Robbins (USA). Due to the new design and operational features that new type of TBM has been named Universal Double Shield TBM with a further extended area of application. Special attention has been given to outstanding torque and thrust capacities combined with the shortest possible length of an innovative shield arrangement, different advance modes in various ground conditions and further developed exploration and treatment facilities.

Jan 1, 2003

By L. R. McCullough

To compete in Today's Market, Mine Managers have two, often conflicting tasks. They are to improve productivity, allowing the mines to compete in our world markets and to improve the environment in which our miners must work. To Maintain and improve the environment to ever increasing acceptable levels requires monitoring. In years passed this required considerable labor, which was provided by the Ventilation Department. They continuously took manual samples using hand-held instruments, sampling gas level such as CO for fire and methane for explosive risks. In addition they took continuous air velocity readings with hand-held instruments which allow them to calculate the ventilation air movement in each area of the Mine. This manual gathering of information was not only expensive and slow but marginal at best. It was marginal because the gas sampling was sporadic rather than continuous, as sample reading taken in one part of the mine would likely not be repeated for another 8 hours. If problems did occur, they likely went unobserved until they reached crisis proportion.

Jan 1, 1990

By Hendri Silaen, Ishak Degei, Nicoll Soumilena, Arwin Edijanto, Avesiena Primadiansyah

PT Freeport Indonesia’s Deep Mill Level Zone (DMLZ) is one of three operating cave mines at the Grasberg mine complex. The DMLZ is situated in the East Ertsberg Skarn System (EESS). It is set 500 m below the operating DOZ mine and located approximately 1,500 m below the surface, making it one of the world’s deepest panel caving operations. Initial undercutting of the DMLZ commenced in the second half of 2015 and has expanded to more than 120,000 m2. High pre-mining stresses and the changes associated with caving, combined with a strong and brittle rock mass, led to significant seismicity and rock bursting. Stress and seismic management are the main challenges for the operation and have evolved significantly from initial undercutting to the present. Considerable amounts of empirical data associated with ground response have been collected, evaluated and combined with historical lessons learned from the Intermediate Ore Zone (IOZ) and Deep Ore Zone (DOZ) mines, which are located within the EESS. This has provided an improved understanding of how the various geological, geotechnical and mine design criteria relate to one another. Additionally, the information has allowed for the development and implementation of improved strategies for stress and seismic management in the DMLZ. This paper describes how these learnings have driven changes in mine geometry, layout, and support systems. Through comprehensive monitoring and detailed inspections, positive changes have been seen in the ground response, leading to an increase in footprint reliability and accelerated cave growth. The paper summarizes the key learnings at the DMLZ regarding stress and seismic management as well as the positive improvements towards sustained and safe caving operations in the DMLZ.

Apr 26, 2022

By Nikhil Dhawan, Shrey Agrawal

Feldspar, a low-alumina, and abundantly available aluminosilicate source, is explored in this study to extract aluminum and potassium values. The sample comprises lamellar layers of microcline and albite with quartz containing 17.4% Al2O3, 9.67% K2O, and 66.3% SiO2. Mechanical (milling) and thermal (alkali and acid baking) activation are investigated, followed by HCl leaching. The milling treatment up to 6 h is inefficient in phase dissociation and results in ~ 40% Al and K dissolution. The alkali thermal treatment yielded 80% Al and 71% K extraction. Direct precipitation of the dissolved Al and Si values yielded alumina purity of ~ 29%, and silica gelation separated SiO2 and improved alumina product (94% purity, 71% recovery). Acid baking yielded 98% pure alumina precipitate; however, Al and K extraction were limited to ~ 45% and 65%, respectively. The microwave-assisted alkali treatment was optimal with 800 W and 10 min processing conditions yielding 80% Al and 71% K recovery with 2.26 kWh/kg alumina of thermal energy consumption. The final product contains γ-alumina with a specific surface area of 87 m2/ g. Keywords Feldspar · Alumina · Potassium ·

May 18, 2022

By B. Yarar

The critical surface tension of wetting (?c) of Cu2S, FeS2, PbS, and MoS2 were determined by plotting flotation recovery versus YLV which was arranged by methanol, and Extra-polating to % R = 0. ?c values ranged from 26 dyne cm-1 for MoS2 to 49 dyne cm-1 for PbS and elemental sulfur gave a value of 31.5 dyne cm-1. Minerals exhibiting natural flotability of different extents, could be separated from their mixtures in the absence of conventional flotation reagents, but controlled ?LV. Correlations between non-stoichiometry and degree of natural hydrophobicity for MOS2 are also reported in this paper.

Jan 1, 1983

By A. H. Srnallbone

Over the past few years the rising production costs, shortage of skilled labor, and increased use of lower grade ore bodies have all added justification for the use of on-stream analysis. A rise in the number of on-stream units indicates that this type of assay is now being accepted as a necessity. On-stream analysis is a much more rapid means of obtaining true assays of a flowing product when compared to wet chemistry or "grab" samples dried, briquetted and assayed by x-ray. In contrast, on-stream analysis uses real time samples assayed every 30 seconds while each stream is assayed every 5 to 15 minutes. It is very important not only to know that a change has occurred in the process, but to know what caused the change and how to correct for the change. For example, in a tailing sample the elements of interest comprise between 1% and 10% of the sample. The remainder is water and gangue. Previously the gangue was not assayed nor were the low atomic number elements used for corrections. In addition to the demand for higher accuracies some concentrators were being penalized for excessive Ca or in some cases 5i in their concentrates. This penalty was being imposed by the smelter in order that they could better cope with pollution problems and prolong the life of their furnaces. But until Ca, 5i, and S are measured on-line, nothing can be done about their control. By accurately measuring these light elements on-line, a complete assay is obtained of the total sample providing improved accuracy for the important products. The justification for complete accuracy assays is now even more obvious because they are possible at very 1ittle increased cost and at no increase in analytical time.

Jan 1, 1977

Arizona copper mines produced nearly 67% of the nation?s copper in 2000, according to a report released by the Arizona Mining Association (AMA). This was a slightly larger share than in 2000 despite large reductions by one major copper producer and production cutbacks at several other facilities. In 1999, Arizona mines produced 64% of the nation?s copper. The report was prepared by economist George F. Leaming of the Western Economic Analysis Center in Marana, AZ. It also showed the copper industry had a $4.6-billion impact on the Arizona economy in 2000, a 26%decline from 1999. This is the lowest financial impact since 1973 and less than half the record reached in 1981of $9.6 billion. The decline is primarily due to a sustained low copper price and a slow-down in the US economy. The mining industry?s impact included combined direct and indirect contributions of $1.283 billion in personal income for Arizona residents, the equivalent of 43,000jobs. More than 5,000 of those jobs were in state and local government and more than 18,000 were in trade and ser-vice businesses. The copper industry directly employed fewer than 8,000 people statewide in 2000. This was down about 23% from 1999.

Jan 1, 2001

By T. W. Martin

In 1973 the percentage of total coal production mined by surface methods in the United States again increased, reflecting a long term trend. Barely three decades ago less than 10 percent of the U.S. coal was produced in this manner. Last year surface coal mines produced 299 million tons, 50.7 percent of our nation's coal. At present there are about 2450 strip mines producing coal in 23 states. While accident rates for surface coal mines are less than underground coal mining rates, surface coal mining is certainly one of the more hazardous occupations. Figure 1 shows the injury rates (fatal and non-fatal accidents) of major U.S. industries. Note that surface coal mining is more hazardous than all other industries except contract construction. Accident rates are about the same as the contract construction industry which uses similar equipment. A limited amount of detailed data was available on elements of surface coal mining jobs and this lack of compiled information made it difficult to assess the significance of the conditions contributing to accidents and to make recommendations that would reduce accident frequency. Therefore as part of the Bureau of Mines continuing research efforts related to the health and safety of the coal miner, a contract was awarded to Theodore Barry and Associates of Los Angeles to conduct an "Industrial Engineering Study of Hazards Associated with Surface Coal Mines."

Jan 1, 1974

By J. Drelich

The flotation limit of fine particles has been re-examined based on the effect of bubble size on contact angle, and a new surface chemistry-limited relationship describing the minimum particle size which can be floated is proposed. In this analysis the critical value of bubble-solid contact radius (re) for which there is no effective attachment between particle and bubble has to be determined. The values of re determined by extrapolation of our experimental data for contact angle versus bubble-solid contact diameter permit the prediction of the flotation limit of methylated quartz particles, with fraction-al coverage of trimethylsilyl groups from 0.07 to 0.51. The limit of fine particle flotation as determined by this theoretical approach agrees well with experimental data reported in the literature.

Jan 1, 1994

By Donald Bishko

Zoning is the legal means by which land use can be controlled. Although zoning laws have historically been quick to prevent quarrying, one is hard pressed to find zoning ordinances which specifically state that land should be used (or held in reserve) for quarrying and prevent other uses such as recreation, housing, etc. This situation is especially true when the mineral to be quarried has a low intrinsic value such as sand, gravel, and crushed stone. The chance of a gold mine surviving urban sprawl is probably considerably greater than that of a sand and gravel pit, even though it is possible that the dollar value of the sand and gravel' would exceed that of the gold mine. The research reported herein was concerned with two major topics: to delineate the real costs associated with the non-protection of our construction aggregate resources (growth minerals) and to develop a computer simulation model which would aid planners in making economically sound decisions. This paper describes the problems faced by the mineral aggregate industry and discusses the role of computer simulation in planning problems. It briefly describes a simulation model which has been developed and previously reported12 and discusses the results of a pilot study conducted on the Tri-City area (Albany, Schenectady, and Troy, New York). 3

Jan 1, 1970

Micro-computers have made it practical to install a high-speed acquisition system to collect data from a longwall face in near real-time. By simultaneous data analyses using neural networks and fuzzy sets, information regarding longwall mining performance and underground behavior can be provided and forecasting longwall instability and detecting geological anomalies is feasible. This paper presents details of the application of neural networks and fuzzy sets in developing on-line evaluation of longwall roof behavior. Techniques for interpreting longwall shield pressures and characterizing roof conditions are discussed.

Jan 1, 1996

By J. David Gunn, Richard W. Lawrence

Gold and silver often occur finely disseminated in sulphide minerals such as pyrite. The use of biological leaching to dissolve pyrite and liberate gold has been extensively studied for a concentrate from the Porgera deposit, Papua New Guinea. The successful operation of a continuous bioleach circuit has been demonstrated. Bioleach residues were cyanide leached to extract gold and silver, and unreacted pyrite recovered for recycle to the bioleach. A 26 day steady state continuous bioleach test on a concentrate assaying 14.3 g/tonne Au and 87.0 g/tonne Ag produced residues with an average 73.5% weight loss and a pyrite oxidation of 88%. Cyanidation and recovery of pyrite from the cyanide residues for recycle gave overall recoveries of 92.1% gold and 74.6% silver. A proportion of the remaining precious metals was found in the bioleach solutions and may be recoverable. Losses to tails were 3.0% gold and 8.1% silver. Final tails represented 14.3% of the feed weight and averaged 1.84 g/tonne Au. A description of the continuous bioleach circuit and bioleach residue handling procedures, and the results of the study are presented. The features and advantages of biological oxidation as a pretreatment for refractory precious metal ores and concentrates are discussed.

Jan 1, 1985

By Augusto Andres Torres Toledo, João Felipe Coimbra Leite Costa, Luciano Nunes Capponi, Diego Machado Marques

Short-term mine planning to open pit requires defining diglines to orientate the excavation geometry. In deposits with multiple chemical variables, it is necessary to have more rigorous control over the feeding of the processing plant. This study analyzes the impact of incorporating multiple chemical variables, the risks generated by geological uncertainty, and the number and starting point locations of shovels in the sequential design of diglines. The objective is to maintain and preserve the low-grade variability, required tonnage, and continuity in sequentially selecting blocks. To achieve this, the probability distribution function of each chemical variable is used while attempting to maintain the ore grades as constant as possible. As it is difficult to maintain multiple elements with low variability simultaneously, the relevance of each chemical variable in the economic benefit is identified. Weights are assigned to signal the importance of each element based on its significance in the recovery process. Several diglines, consisting of sequential blocks, were generated on the same seed point to meet the weekly ore demand. Optimizing the stationarity of multiple chemical variables, accessing and reducing the geological uncertainty performed through geostatistical simulation of grades and transfers functions, considering the number and location of shovels, while meeting the operational restrictions of the selected blocks. The optimized digline with the multiple grades as stationary as possible is retained. The process is illustrated in a phosphate mine. The results show that the proposed method is viable and allows for a quick solution.

Jun 25, 2023

By N. Arbiter

Introduction The theoretical analysis of tumbling mill energetics and performance has largely neglected mill dimensionality, and, in particular, the importance of the length/diameter (L/D) ratio. This is in spite of the fact that practice varies substantially: geographically, for semi-autogenous and autogenous mills, as between North America and South Africa/Scandinavia, and historically, for overflow ball mills, for which the L/D ratio has increased significantly from the earliest small mills to the largest mills currently. The present study is concerned primarily with the influence of the L/D ratio on the design and operation of overflow ball mills, on the occurrence of the overload phenomenon, and on the limits, if any, it may impose on mill capacities because of critical pulp axial velocity limits. It will be shown that the shape factor is of major importance in this area and that its adjustment to the extent that this is practical should remove the diameter limitations previously postulated for this mill type. Dimensionality in mill design Ball mill shape factors in the period prior to 1927 (Taggart, 1927) averaged 1.1/1 for 29 center discharge mills and 1.0/1 for 30 peripheral discharge mills. With the resumption of new plant construction after the 1930s depression, the Morenci concentrator continued the 1/1 ratio with its 3.1 x 3.1 m (10 x 10 ft) mills. The ratio was increased progressively from then on, reaching 1.6 and 1.8/1 for the largest overflow mills currently. As shown recently (Arbiter, 1989), this is in sharp contrast to autogenous and SAG mill shapes, for which the ratio averages 0.4 in North America. On the other hand, South African practice, starting at the turn of the century with autogenous mills having 4/1 ratios, moved toward 1/1 until recently, when a 2.5/ 1 ratio mill was installed. The reasons given for such divergent practice for mill shape factors are in some respects contradictory and generally inconclusive. The most complete discussion from a practical viewpoint (Dor and Bassarear, 1982) is limited to primary SAG and autogenous mills. Considerations of ball mill dimensionality have had a twofold direction. On the one hand, it has been argued that ball mill efficiencies should increase with increasing diameter and that the specific energy for a particular grind should be reduced accordingly. An inadvertent test of this idea at the Bougainville operation (Burns and Erskine, 1983) resulted in drastic underpowering, which led to failure to reach design capacity until additional mills were installed. This can be taken as strong evidence against any increase in efficiency with diameter. In another direction, it has been argued (Arbiter and Harris, 1982, 1983) that there is a limit to ball mill diameters because of the demonstrable limit to axial flow velocities evident in the overload phenomenon, which as a fact is incontrovertible. But that it places a limit on mill diameters overlooks the evidence given below that appropriate variation in the L/D ratio will permit major increases in diameter, limited only by constructional or economic factors. The present study was directed toward quantifying the overload phenomenon through examination of the influence of mill dimensionality variation and mill operating variables on its occurrence. It is shown that varying the operating conditions, specifically the load fraction and the fraction critical speed, can reduce the risk of overload for existing operations; while appropriate decreases in the L/D ratio can minimize the risk in the design of new circuits. Ball mill overload Ball mill overload is a consequence of the approach to a critical velocity with increasing feed rates or circulating loads. Although the effect has been known for over 50 years, there have been no previous attempts to quantify it. The following description of the ball mill as a flow system is the preliminary to a quantitative analysis: 1) In the absence of a ball load, axial flow of pulp through a mill resembles open channel flow, except for disturbances near the shell due to shell/lifter rotation. 2) For a given ball load (Lf), the void fraction available for pulp hold-up (H) for the ascending portion of the load is approximately 0.4 Lf. In the descending portion, it is greater than this and increases with fraction critical speed (Fc) because of load expansion. 3) Increasing the feed rate increases pulp hold-up and progressively fills the available void space. At a critical flow rate, which depends on system geometry, hydraulic head and pulp rheology, void filling reaches its limit; a pool forms rapidly and fills available space outside the ball load and up to the overflow level. Prior to this, pulp discharges mainly along the ascending rim of the overflow. 4) For a small mill (Lo et al., 1990), it has been shown that with increasing feed rates the critical filling is at or near 50% of the mill volume. Power drops rapidly when this level is reached, as required by the torque formula. 5) The transition to overload is associated with the following phenomena: a) The decrease in power draw. b) Damping of mill sound. c) Reduced comminution of coarser feed sizes, probably due to reduced direct impact in the presence of a pool. d) Increased circulating loads, which further intensify the overload, and increased density of cyclone underflows, which can lead to roping. e) Conditions beyond overload are not known because feed rates are not delibelrately increased beyond this point. f) The existence of the phenomenon limits the capacity of a mill with a fixed set of operating conditions and can prevent the balancing of hard ore feed rate decreases by increases in soft ore rates.

Jan 1, 1992

By R. S. Conti, M. R. Krump, C. P. Lazzara Ph. D., R. L. Derick, L. L. Chasko

A mine fire is one of the most challenging safety issues facing a mine operator and can occur at any location underground. The National Institute for Occupational Safety and Health (NIOSH) and the Twentymile Coal Mine, CO, conducted in-mine experiments to determine the capability of high-expansion foam for addressing underground fires. The tests were conducted in sloping entries with high-capacity diesel-powered foam generators. The following results were observed: a well-designed, maintained and properly operated high-expansion foam generator can propagate a foam plug hundreds of feet in steep upward-sloping (20%) multiple entries against a ventilation pressure; stoppings and partitions designed to contain the foam plug in upward-sloping entries must be substantially constructed; and the predicted quantity of foam concentrate can be significantly less than the actual amount used, due to foam losses, such as bubble breakage on dry surfaces. This paper describes the production of high-expansion firefighting foam and discusses the in-mine experimental tests and results.

Jan 1, 2010

By W. Fletcher

Startup of a new plant is a very interesting time. It generally is also a very demanding and often frustrating period. Startup problems can become serious ii they impede the early achievement of full plant production. Raw materials are wasted, manpower is unproductive, field modifications can be extremely expensive and market commitments may not be met. This paper outlines a well proven and highly successful approach to eliminating many of the problems and frustrations often experienced in plant startup. The procedures and methods summarized in the following paragraphs have been developed over a period of 25 years during which some 90 plants were started up.

Jan 1, 1981

By J. F. Burr, D. Snyder, W. Calder

"INTRODUCTION “Explosion protection” refers to techniques used to minimize the potential for electrical and electronic equipment to create an ignition while operating in a hazardous location (HAZLOC). In particular, U.S. coal mines are required to use equipment in certain areas of the mine that has been approved by the Mine Safety and Health Administration (MSHA) for use in a methane and coal dust environment to limit the risk of the equipment creating an ignition-capable spark or a thermal energy ignition. In general, MSHA’s regulations for explosion protection recognize two techniques: the use of explosion-proof enclosures (XP boxes) and 2-fault intrinsic safety for electrical and electronic equipment. For 2-fault intrinsically safe (IS) equipment, the U.S. mining industry requirements are unique in that the equipment is certified by MSHA. The criteria MSHA uses to evaluate equipment for intrinsic safety, as required by Title 30 of the Code of Federal Regulations (30 CFR) [30 CFR, Sections18.20(b), 18.68(a)(1), 19.1(b), 20.1(c)(2), 22.6, 23.6, and 27.20(a)], is published in a document referred to as ACRI2001 (1). Outside of U.S. underground mining, many industries and countries accept equipment that is designed to a consensus-based international standard (ANSI/ISA 60079-11 is the U.S. version). The purpose of this study is to provide an overall assessment of the ANSI/ISA 60079-11 standard on 2-fault intrinsic safety and the MSHA ACRI2001 acceptance criteria, and determine if the ANSI/ISA document can be an alternative to ACRI2001 while maintaining an equivalent or better level of safety for miners. There are many drivers for this study. First, immediately after the 2006 MINER Act (2) was passed, there was a perception in the industry that the unique MSHA requirements were contributing to delays or restricting the availability of equipment to meet the requirements of the Act. Additionally, the MINER Act required that mine operators submit an accident response plan for MSHA approval, to be reviewed every six months. MSHA is required to determine if these plans reflect the most recent credible scientific research and make use of currently commercially available technology. There is a concern, as expressed in a recent National Academy of Science (NAS) study (3), that the unique requirements for U.S. mining may be limiting the technologies available to coal miners and technology options for mine operators. Accordingly, the study recommended that the National Institute for Occupational Safety and Health (NIOSH) and MSHA re-examine their technology approval and certification processes to ensure they are not deterring innovation, and explore opportunities to cooperate with other international approval organizations to harmonize U.S. and international standards without compromising safety. Lastly, the National Technology Transfer Act (NTTA) and the accompanying Office of Management and Budget’s OMB-119 circular mandate the use of consensus-based standards by federal agencies. However, for U.S. mining operations, MSHA cannot accept such standards unless a determination is made that such standards provide an equivalent level of protection for the miner, per the requirements of the 1977 Mine Act, the 2006 MINER Act, and 30 CFR Part 18 regulations."

Jan 1, 2017

By P. Raffinot, C. H. G. Bushell, S. Hodgson, E. F. Milner, H. J. Bosshardt, N. Draper, D. E. Pickett, Edwin T. McKnight, J. Prieto

Zinc ranks third and lead fourth in world nonferrous metal conumption. In 1972 the free world consumed 3,333,000 mt of refined lead and 4,422,000 mt of refined zinc. The lead tonnage was divided into the following end uses: storage batteries (42%), lead allows (7%), pigments and chemicals (11%), cable sheathing (9%), gasoline additives (13%), pipes and sheets (8%), and miscellaneous uses (10%). The corresponding distribution of zinc uses in 1972 was galvanizing (39%), die casting (26%), brass (15%), rolled zinc (8%), oxide excluding American Process zinc made directly from concentrate (6%), and other uses (6%). The pattern of usage of lead and zinc varies considerably from country to country. During 1972 in the USA 49% of refined lead was consumed in storage batteries, 19% in gasoline additives, and only 3% in cable sheathing, whereas in the United Kingdom 28% of the lead was consumed in batteries, 14% in cable sheathing, and 14% in gasoline additives. For zinc the major US uses were die casting (41%) and galvanizing (37%), with about 3% consumed as rolled zinc; corresponding usages in the same applications in France were 12, 28 and 28% and in Australia 12, 67, and 3%, respectively. The outstanding resistance of lead to corrosion, its density, and

Jan 1, 1985

By Robert A. Jankowski, J. Drew Potts

The Bureau has identified several ventilation techniques and a water spray application for reducing auger mining dust levels. The dust control methods all work by improving air quantity or flow patterns at the face. Section airflow is maximized by reducing air leakage between entries. Sealing stoppings with brushed- on mortar, containing fiberglass additives, reduces leakage. Heading airflow is maximized by properly installing the brattice curtain and keeping the area between the curtain and rib free from obstructions. The brattice curtains should be hung at least 1.7 m (5.5 ft) from the ribs in 7.9 m (26 ft) wide entries and kept less than 4.6 m (15 ft) from the face. A tight curtain seal is created by using the mine air pressure to force the curtain against the timbers, roof and floor of the mine. Using both a blowing curtain and an exhausting curtain will isolate some face workers from dusty air. The Bureau designed, auger clearer spray system, can be used to force heading air across the active face. Laboratory tests indicate that the auger clearer spray system can increase face velocities by more than 0.5 m/sec (100 ft/min), while using less than 0.13 l/sec (2.0 gal/min) of water.

Jan 1, 1989

By T. I. Martinovic

A major investigative' program was conducted on the development of an acceptable substitute organic binder for bentonite for the pelletizing of the Iron Ore Company of Canada's Carol concentrate. The first phase of the program covered an evaluation of several organic binders in laboratory tire balling in order to identify the most promising potential organic binders. Based on the green ball quality obtained, Peridur C-10, Alchem 8C26 and peat moss were selected for full scale pot-grate testwork. The selected binders were tested with various pelletizing mixes (acid, dolomite, limestone and fully fluxed). The pot-grate test results showed that an acceptable pellet quality can be produced with the use of Peridur C-10.

Jan 1, 1989