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By M. D. Johnson

The San Andres mine, located in the western mountain region of Honduras, Central America, contains epithermal gold mineralization as disseminations, in quartz-chalcedonic veins and stockworks. Host rocks include a thick sequence of Cretaceous to Tertiary conglomerates, sandstones, rhyodacite tuffs and andesites. Mineralization is controlled by northwest and northeast striking, normal faults and fractures which dip steeply east and west. Currently, 500 STPD of 1. 93 Grams/MT gold ore is being open-pit mined from the conglomerate and andesite. The mine is at the west end of a mineralized zone that can be traced over an area of approximately 1.5 square kilometers.

Jan 1, 1993

By D. Jacobson

There are many different machine design parameters and variables that affect the production of cone crushers in a mineral processing plant. One of these variables is the speed at which the cone gyrates. In June 2012, pilot tests were conducted to determine the production and operating conditions over a wide spectrum of speeds. mining operations will have variations in the ore characteristics and feed rates that result in non optimal use of the crushing energy and inconsistent process control. This study documents the measured variation in production and energy efficiency with two operating modes. Overall, the results match the theory, and the capacity and power fluctuated by 25% and 17%, respectively, when operated at full capacity, with an overall trend of slightly coarser discharge size distribution as the speed decreased. Using a fixed feed rate, the crushers had 4-14% higher production when operating at the optimal speed. Secondary analysis showed that the power draw consistency and magnitude of power spikes could also be improved by optimizing the cone speed. Varying the speed of the crusher can be used to tune the crusher during commissioning, adapt to changing ore conditions over long periods of time or be used as a dynamic input into an advanced control system. each action above would serve to make the crusher(s) more productive, more energy efficient, impove process control and would also have mechanical benefits for the machine

Dec 1, 2013

By John Bley

White Pine Copper Company operates a mine in the Upper Peninsula of Michigan using a modified room-and-pillar system for extracting ore.1 Mine production averages 15, 500 tons per day. The terrain is predominantly level and the depth of cover varies from 500 to 800 feet over the present workings with depth increasing as mining progresses down the dip. The deposit is situated on a low angle plunging, anticlinal structure composed of sedimentary beds. Local terminology for the stratigraphic units which contain the ore horizons, named in ascending order, is as follows: lower sandstone, parting shale, upper sandstone and upper shale. The ore zones of copper enrichment are, contained in the parting and upper shales. The parting shale will vary; from 6. 5 to 9.0 feet thick; the upper sandstone from 2 to 10 feet; and the upper shale from 1. 5 to 6. 5 feet. Two variation of the room-and-pillar system are used to extract the ore depending on grade and local ground conditions. One system is called the parting shale method and recovers the parting shale only. This system requires a mining height of 8 1/2 feet to accommodate large-scale, high-capacity equipment and this working height is obtained by removing approximately 1 to 2 feet of the barren lower sandstone with the ore. Under the second system, locally termed full column, the mining height is about 14-1/2 feet and this system removes the parting shale, the upper sandstone and the lower mineralized zone, approximately 1-1/2 feet, of the upper shale at which point it is possible to maintain a competent back (Fig. 1).1

Jan 1, 1964

By D. A. Rice

In pelletizing magnetic iron ore concentrates, about 6 to 18 lb of bentonite binder are added per ton of concentrate during balling to give green and dry strengths to the unfired pellets. The dry bentonite is mixed into the filtercake following the magnetic concentration circuit of the plant. The filtercake contains about 10% moisture, and the water may be of highly variable composition, depending on the particular concentration circuit, its source, and the ore being treated. A knowledge of the parameters that influence the balling operation is essential for optimal efficiency in producing balls strong enough to survive the treatment during transportation from the balling device to the furnace and during the firing operation. Furthermore, very weak unfired pellets probably contribute to high dust losses and poor furnace control. The factors that determine the strength of unfired pellets have received considerable study (1,2,3,4). Tigerschiold and Ilmoni(l) pointed out the significance of the surface area of the particles being pelletized. Merklin and Childs(2) reported on the effect of material size consist, chemical additives such as soda ash, lime, balling time, ball size, and the moisture content of green balls. Iwasaki and Zetterstrom(3) examined the detrimental effects of fatty acids on the pelletizing process. Stone(4) correlated the zeta potential of bentonites with the strength of unfired pellets. He found that, in general, a high negative charge on the bentonite is conducive to good balling. Although studies of pelletizing in the iron ore industry deal with a complex multicomponent system consisting mainly of the iron mineral, bentonite, and water, the last of these has received little attention, particularly with regard to the possible effects of mill water chemistry on the pelletizing process and the strength of unfired pellets. It is widely known that the composition of mill process water is an important variable in unit operations such as flotation, thickening, and filtering and that impure water can lead to poor metallurgical results. The results of the present study indicate that the chemistry of the mill water may be an important variable in the balling operation as well. Knowledge about this variable should contribute to lowering bentonite and oil consumption and to improving furnace performance. Such knowledge can also assist those who use balling in evaluating commercial bentonites for plant use as well as those investigating balling from a more fundamental point of view In this study we have examined 1) the effects of replacing mill process water in the magnetic concentrate filtercake with pure water, and 2) the effects of several chemical impurities that are commonly found in mill process waters upon the strength of unfired pellets. The results are interpreted in terms of the zeta potential of the bentonite and magnetite particles. The data lend support to a suggested mechanism whereby the electrostatic attraction of oppositely charged particle surfaces causes bentonite-bentonite and magnetite-bentonite bonding. This mechanism would also explain why several chemical additives e.g., sodium carbonates and phosphates, have such marked effects on increasing the strength of unfired pellets.

Jan 1, 1971

By Kiran B. Chaudhuri

The Kivcet process is technologically the most advanced of the new lead smelting processes which have been developed over the past decade construction of the first commercial scale Kivcet lead smelter which has been designed by KHD Humboldt Wedag for Sociedad Complejo Metalurgico de Karachipampa is at present nearing completion in Bolivia. Some of the measures taken to minimise emission of harmful elements as well as methods employed to minimise energy consumption in the design of the smelter, are presented. Results of recent testwork which have added to the flexibility of the process are summarised.

Jan 1, 1982

By Bo-Hyun Kim, Mark K. Larson

"Bumps remain a critical safety concern in underground coal mining, despite significant advancements in engineering controls. These events occur when relatively high stresses in a coal pillar in conjunction with low confinement exceed the pillar’s critical capacity, causing the pillar to rupture without warning. Floor failure often occurs and can be an indicator of bump risk, as it is strongly associated with brittle failure known as spalling under a lower confinement. Significant floor heave at a western U.S. coal mine has been observed where the pillar width in a three-entry gate road system was 53 m, while no significant floor heave was observed where the pillar width in another gate road system was 22 m. This phenomenon was widely reported in the region and at another mine in another western coal field it has been reported that larger pillars are always associated with floor heave. In this study, a spalling model based on the S-Shaped brittle failure criterion and associated with the anisotropic characteristics of the materials was used in FLAC3D. The observed floor heave phenomenon associated with the larger pillars was demonstrated using this modeling approach.INTRODUCTIONFloor failure in a coal mines often inhibits longwall mining operation, as large displacement of floor strata, known as floor heave, interferes with travel, access, ventilation, and equipment operation. Aggson (USBM 1978) investigated floor heave in West Virginia where underclay is relatively strong and high horizontal stress is present. It was reported that the failure mode is mainly by buckling rather than by plastic flow.This study was developed as part of an effort by the National Institute for Occupational Safety and Health (NIOSH) to better understand rock mass behavior in longwall coal mines in highly stressed, bump-prone ground. More specifically, this study attempts to explain through numerical models small energy releases in the form of floor heave. The authors believe that brittle rock failure is associated with both floor heave and some bump mechanisms, if not directly involved in their cause."

Jan 1, 2019

By John T. Awald

The efficiency of search for porphyry copper and molybdenum deposits can be greatly increased. A model exploration system is here presented to show how geologic literature data and field surveys can be aided by modern statistical and computer methods. Assuming that major ore deposits occur in clusters, it is possible to evaluate rapidly large areas at modest cost and with minimum investment risk. An example of the success of these methods is given. Clusters of mine density and metal production covering the State of Nevada were plotted and machine contoured. Selection of a 47,300 square-mile area on the Nevada-California border was made. Within this area, detailed analyses were performed, and a 3000 square-mile site was picked for subsequent field examination. To initiate efficiently the field work, 81 stream sediment geochemical samples were collected and analyzed for copper, zinc, and molybdenum. These data established local variance and necessary statistical controls. Following this, a total of 151 geochemical samples were used to survey the 3000 square-mile area. Computer contouring of results directed attention to specific target areas for detailed geologic analysis, property acquisition and drilling. A total of 22 copper-zinc and molybdenum anomalies were identified. Four of these were recommended for further exploration. To date, two have been acquired and drilled and preliminary ore reserve and financial analysis have been made on one. The estimated cost and time for the regional exploration phases leading to these targets were thirty man-weeks and $16,000.

Jan 1, 1973

By K. F. Clark

Some 30 years ago Stevenson (1940) underscored the fact that most of the world's molybdenite production came from the Cordillera that comprise the western part of North America, and that the northern extension of the Cordillera into British Columbia is not essentially dissimilar to the United States section. Therefore, in view of the successful development of molybdenite resources in the United States, he concluded that prospecting in the Canadian section f the Cordillera should be encouraged. However, as recently as seven years ago Schneider (1963) remarks that no large economic deposits have been brought into production in British Columbia, although one important property was being evaluated. World production f molybdenum ore was sporadic until 1900. Among the very early producing areas in North America were Arizona and New Mexico in the United States. World War I provided the stimulus that brought the Climax deposit into production, and in 1920 Questa, New Mexico followed, although the ore came from fissure vein deposits at that time. In the early part f this century high grade deposits in various parts f the world containing 5-20 percent MoS2 were exploited. In Canada most production during World War I came from Quebec and Ontario. With the build-up of world tension and the inception f hostilities in 1939, British demand was satisfied by United States producers (Schneider, 1963).

Jan 1, 1970

By E. R. Phelps

It is very seldom that a completely new project is initiated. Most projects consist of a new and different combination of proven techniques or a combination of proven and new techniques. The theme of my discussion today falls in the second category. Proven methods of water transportation and proven methods of bulk material handling were combined with some new techniques which resulted in a new market for the coal industry and a cheaper fuel supply for the power industry. In this respect I refer to the water transportation of coal from the West Kentucky coal fields to coal burning power plants on the west coast of Florida. Peabody Coal Company, Florida Power Company, Gulf Coast Transit, and Mid-South Towing combined their individual requirements and "know how' to make this project possible. I feel the results speak for themselves since it proves that coal can be mined in the West Kentucky coal fields, transported some 1,200 miles by river barge and some 450 miles by ocean going barges, be delivered to power plants on the west coast of Florida and still meet the competitive prices of gas and residual oil. The unique parts of this project concerns the storage and trans loading facilities for coal from river equipment to ocean equipment near New Orleans; the return haul of phosphate and it's trans loading from ocean to river equipment at the same facility and the use of large ocean going barges. In detail it works as follows: Coal is mined in West Kentucky and loaded into 1,500 ton river barges on the Green River. It is then transported some 100 miles to the mouth of the Green on the Ohio River, then 200 miles down the Ohio to it's convergence with the Mississippi, and then 900 miles down the Mississippi to the trans loading facility some 40 miles South of New Orleans. There it is transferred from the 1,500 ton river barges to 18,000 ton ocean going barges and hauled some 450 miles across the Gulf of Mexico to the Gannon Station of Tampa Electric at Tampa entailing a total haul of approximately 1,650 miles.

Jan 1, 1966

By K. J. Shoop

Coal-fired power plants in the U. S. produce approximately 20 million metric tons of Hue-gas desulfurization sludge every year. The sludge mainly consists of calcium sulfite and calcium sulfate. This can be converted to synthetic gypsum, which in principle could be a marketable commodity. However, many Hue-gas scrubbers produce sludges that do not meet the quality requirements of the gypsum industry, and so their sludges must be landfilled as a waste. If a suitable method can be developed for removing certain objectionable impurities from the sludge, it will then be possible to market the material instead of landfilling it. A sludge from a natural-oxidation scrubber was characterized, and the main contaminant was found to be calcium carbonate from unreacted limestone. Based on the particle size and the characteristics of the material, froth flotation was considered as a method for purifying the material. Since no flotation studies have ever been carried out to separate calcium carbonate particles from calcium sulfite or calcium sulfate particles, it was first necessary to select a suitable collecting reagent. Based on Zeta-potential studies, the most suitable reagent was found to be an octadecylamine-octadecylguanidine complex. Results are presented for flotation of calcium carbonate from scrubber sludge using this reagent, in a reverse-flotation process. The best results were obtained when the sludge was first pre-cleaned using a water-only cyclone to remove unfloatable coarse limestone, followed by froth flotation to remove fine limestone.

Jan 1, 1998

By H. R. Green

The design of slurry launders has usually been based on strictly empirical concepts. An examination of the most common procedures reveals that they do not account for many of the variables that are recognized as Significant for slurry transport. These may include flow rate, volume concentration of solids, solids specific gravity, solids size distribution particle shape, launder geometry, and roughness of the wetted surface. It was decided to develop a design procedure, which would accomplish two things: First: Take into consideration most of the known significant variables and systematize the procedure to assure consistent results. Second: Provide a rational basis for examining and utilizing operating data to refine and improve the system. To accomplish this, it was necessary to develop a basic design concept. This concept has been developed through a complex development history, and yet still appears workable and technically sound. That concept may be outlined as follows: First: The solids transport velocity is the fundamental basis for slurry launder design. Second: The stream configuration, hence the launder size, is integrated with the solids transport velocity so that the actual stream velocity exceeds the solids transport velocity. Third: The launder slope is that which will achieve the required actual stream velocity. It was originally planned to utilize a series of charts and nomographs for the design procedure. A nomograph, based on data given in Taggart's Handbook of Mineral Dressing (1), was developed to determine the solids transport velocity. The velocity thus determined was used with Manning charts to determine the launder slope .and configuration as though the fluid were water. It was planned to utilize a "slope adjustment factor" to increase the slope to compensate for the apparent viscosity of the slurry. All attempts to develop a suitable correction factor failed. This failure was principally due to the complex relationship of viscosity to the required slope. It was then decided to revise the entire approach. Specifically, the present procedure is based on the Darcy-Weisbach flow equation, which is rationally preferable to the Manning equation. The Camp (2) equation is used for the solids transport velocity. The necessity of reference to the Moody curves for the Darcy friction factor is avoided by using the Colebrook and White (3) equation. The resulting mathematics require iterative solutions for several of the unknowns.

Jan 1, 1978

By Manchao He, Xiaoding Xu, Yubing Gao

Based on the geological conditions of the 15,600 working face in Pingba Mine, this paper establishes the mechanical model of the stress distribution in the pre-working face under the influence of mining pressure based on a hemispherical stress distribution theory. The analytical formula for the stress distribution is determined. A numerical model for the mining pressure under the influence of key layer is also constructed, and dynamic stress interruptions under various roof-cutting relief schemes in different horizontal positions are simulated. The stress distribution of the stope under different pressure relief schemes is analyzed and compared with the theoretical results. Results show that blasting roof cutting blocks the transmission of the abutment pressure to the working surface to the coal and rock mass in front of the working face, reducing the significant working face mining pressure influence range, which decreases as blasting pressure relief position moves toward the working face. Stress accumulation in the coal in front of the working face also increases as the blasting pressure relief position moves toward the working surface. Engineering practice results show that blasting cutting can effectively block the propagation of mining pressure, and the stress impact range is shortened by approximately 1/3.

Jul 31, 2021

By Ismail Adeniyi Okewale, Hendrik Grobler

Significant factors believed to be responsible for tailings dam failure which can lead to loss of lives, destruction of properties and assets or damage to the environment are the mechanics of material characteristics and static liquefaction. It is therefore essential to understand the tendency for the likely occurrence of liquefaction in tailings dam as well as mechanics behind the behaviour of this tailing materials (inevitable products of ore mining, mineral and metallurgical processes). This study aims to present the mechanical behaviour and susceptibility to liquefaction of a gold tailings sample. Series of oedometer tests were conducted to determine the compression behaviour and the possibility of transitional behaviour or not. Index and physical tests were conducted to determine engineering grading descriptors which was directly linked to compression behaviour and comparisons made with other tailings. Microstructure was studied to understand the effects of fabric on the behaviour of samples and the chemical composition and mineralogy were also investigated. The compression paths slowly converge and the gold tailings might be said to be less transitional. The gold tailing is less dependent on initial state. The loosest gold tailing is more susceptible to liquefaction and compared to other gold tailings, this sample is less susceptible to liquefaction.

Apr 19, 2024

By Hugh D. Olmstead

Typical operations necessary to develop oil and gas reservoirs are reviewed, including exploration, drilling, cementing, completion, work over, sand control, stimulation, enhanced recovery and production. The role of industrial minerals in these operations is described. Some of the opportunities and problems affecting increased use of industrial minerals in drilling and stimulation are discussed.

Jan 1, 1983

By Daniel A. Wasson

A review of the history of boosters is given with examples of the prominent types that have been made commercially. The performance and design features of protected core boosters are discussed in detail. A new gel booster is discussed with its advantages to the mining industry. Booster size, hole configuration, performance and safety aspects are reviewed. Important features of boostering in above ground and underground blasting are related to an economic analysis. This presentation will cover different types of boosters commonly used in mining. However, for the purpose of this paper a few terms need to be defined to clarify the action of different parts of a priming system. Booster* - "Any unit of explosive or blasting agent to be used in a charged hole for the purpose of perpetuating or intensifying the initial detonation". Primer* - "A unit, package or cartridge' of explosives containing a detonator or detonating cord which is used to transmit detonation to other explosives or blasting agents". Detonator - A device for initiating a detonation from a flame or a small electric current, the former being represented by a fuse cap and the latter an electric blasting cap.

Jan 1, 1981

By Y. Lahaye, O. Martinsson, B. I. Palsson

"Four main and three minor rare-earth-element (REE)-bearing minerals were identified and quantified in the Kiirunavaara apatite iron ore tailings using optical microscopy, an electron probe microanalyzer (EPMA) and a mineral liberation analyzer, and their chemical compositions were analyzed by the EPMA and laser ablation inductively coupled plasmamass spectrometry. REEs are shown to be contained in the minerals apatite, monazite, allanite, titanite, zircon, thorite and synchysite. In zircon, thorite and synchysite, they occurred in only trace amounts and contributed limited amounts to the total REE budget, and these are consequently of minor importance. Monazite occurred as inclusions in apatite and as free particles, 90 percent liberated. Allanite occurred to some degree in mixed grains with magnetite but also as free particles. Monazite mainly reported to the apatite concentrate, while allanite and titanite largely went to the tailings, the latter preferably to those fractions smaller than 38 µm. The amount of titanite in the finest tailings fraction was 2.3 weight percent, containing close to 1 percent REEs, with heavy rare earth elements (HREEs) making up 28 percent of the total REEs. However, a texturally distinct group of titanite grains showed an HREE/REE ratio of up to 67 percent. Furthermore, titanum dioxide analyses indicate that titanite is preferentially released into the tailings from the secondary magnetic separation step in the concentrator. Our data therefore suggest that titanite, occasionally enriched in HREEs, can be extracted from the processing stream and might thus be considered a new source for REEs at Kiirunavaara and similar deposits.IntroductionDeposits of apatite iron ores are the major source for iron in the European Union and are exploited by Luossavaara-Kiirunavaara AB (LKAB) in several mines in the Kiruna-Malmberget area in northern Sweden. Kiirunavaara is the largest of the apatite iron orebodies in Sweden, comprising more than 2 Gt of iron ore with 60 to 68 percent iron (Fe). At the end of 2015, the total accumulated production of crude ore from openpit and underground mining since the start of mining was 1,566 Mt, with ore reserves of 644 Mt grading 46.3 percent Fe (LKAB, 2016). The deposit defines a steeply dipping tabular body, approximately 6 km long and up to 180 m thick, extending at least 1,500 m below the surface. It is situated in the middle part of the Palaeoproterozoic Kiirunavaara Group of volcanic rocks and follows the contact between a thick pile of trachyandesitic lava and overlying pyroclastics (Martinsson, 2004)."

Jan 1, 2017

By James L. Evers

The development of hydraulic mining in the United States has had an uneven record of progress. The shifting of coal extraction from one government agency to another has not provided an environment for a focused national research program. This paper presents an update on hydraulic mining research in the United States. A comparison is made between the development of the technology in this country and the state of the art in other parts of the world.

Jan 1, 1983

The Bureau of Mines, U.S. Department of Interior, is currently investigating strata interactions associated with the mining of multiple coalbeds. Strata interactions due to subsidence are a result of undermining. This study involves underground investigations at a mine where room and pillar panels are actively being developed through an area of the Sewickley Coalbed which was subsided by undermining of the Pittsburgh Coalbed. Research to date indicates that undermining had little effect on upper mine pillar stability, but had a more severe effect on the development and maintenance of entries. Roof-to-floor measurements recorded over four times more convergence in entries developed over gob as compared with entries developed over support pillars in the lower mine. Consequently, this has led to major roof falls in the active panels, resulting in considerable production delays and supplemental roof support costs.

Jan 1, 1985

By D. R. Forshey

Although the United States has the world's largest coal reserves, most of those reserves will require underground mining. Unless underground injury and fatality rates are reduced, the number of disabling injuries and accidental deaths will grow as coal production increases. The Bureau's Mining Research Program, designed to reduce the exposure of workers to hazards while improving productivity, is described. Major program thrusts are 1) re- moving the hazards or removing miners from the areas in which hazards exist, 2) developing ways to improve management decisions and mining skills so as to reduce the incidence of situations requiring human involvement in hazardous, emergency, or corrective tasks, and 3) improving, or at least maintaining, productivity so that both the tangible (dollar) and intangible (human) costs are reduced. In these ways, projected demands for coal can be met using resources without an increase in human suffering.

Jan 1, 1987

By R. L. Thoms

Compressed air energy storage (CAES) has been initiated recently in West Germany with a 290 MW peaking-power facility utilizing twin cavities with volumes of 150,000 m3 each in the Huntorf salt dome (1). The implementation of CAES in several geologic media is being studied carefully in the United States (2); however, the use of cavities for CAES in salt domes or salt anticlines and the associated long-term stability of these cavities is the specific topic of this paper (3). The cavities represent an essential element in a CAES plant, and thus should be designed and operated to perform with "stability" over the planned life of the overall facility. As employed herein, cavity stability implies the extent to which an acceptable amount of cavity storage volume can be utilized with economic maintenance over the planned life of the overall facility. In this sense stability is relative to the planned use of a storage cavity. Figure 1. depicts the salt deposits of the United States (U.S.) (4). The Gulf Coast region includes the only confirmed salt dome basin in the U.S., although salt structures also exist as anticlines in the Paradox Basin of Utah and Colorado. Because of the relatively small amount of elevation change within the Gulf Coast, CAES in salt dome cavities appears to hold considerably more potential for energy storage in this region than more traditional "pumped hydro" systems.

Jan 1, 1979