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By M. K. Mohanty

A modified coal flotation characterization procedure, referred to as the Advanced Flotation Washability (AFW) technique, has been developed which, compared to traditional procedures, provides a more accurate prediction of the optimum separation performance achievable by a froth flotation process. The AFW procedure uses a vertically-baffled flotation column which provides enhanced selectivity among particles of varying degrees of hydrophobicity due to a plug-flow environment and an improved carrying-capacity environment. In fact, the separation performance of the traditional procedures was enhanced by increasing the possibility of operating under improved selectivity conditions. The separation performance provided by the AFW procedure was superior to that obtained from multiple cleaning stages using commercially-available flotation columns under both kinetic and carrying-capacity limiting conditions.

Jan 1, 1997

By J. Hanna, I. Anazia

Selective fatty acid flotation of carbonate gangue from sedimentary apatites was achieved under acidic pHs without a specific apatite depressant and without conditioning the pulp with the collector. Concentrates (35% to 38% P2O5, 0.8% to 1% MgO) with high recoveries (80% to 85%) were produced from Indian and Chinese dolomitic ores. The calcareous and/or dolomitic-siliceous phosphate ores of South Florida and Utah required two flotation stages: one for carbonate removal, as mentioned above, and the other for phosphate recovery from siliceous gangue. Preliminary tests on the South Florida ore gave concentrates analyzing 29% P2O5 and 0.8% MgO, with a P2O5 recovery of 76%. A concentrate obtained from the western US phosphate ore from Utah assayed 30% P2O5 and about 1% MgO, with a P2O5 recovery of 89%.

Jan 1, 1988

By Mahmud Esad Uluer, Seyed Hassan Amini, Aaron Noble, Matt Shigo

Airborne coal mine dust is an unintended and unavoidable result of coal extraction operations that raises notable health and safety concerns. Modern mining techniques utilize several dust mitigation strategies, and the flooded-bed dust scrubber (FBS) is one such technology used extensively on continuous miners. The conventional static panel filter, instrumental in most scrubber designs, is fundamentally limited in collection efficiencies due to a high clogging rate and a tradeoff between mesh density and airflow rate. To overcome these deficiencies, a novel vibrating mesh panel is proposed. In this study, a laboratory-scale dust scrubber unit was used to investigate the efficacy of concept as compared to that of a static mesh. A statistical design of experiments study was first used to determine the effective vibrational parameters and scrubber operational parameters on dust collection and clogging mitigation. Optimized results from this study were then evaluated against those of a static mesh to determine performance improvement while investigating the mechanisms controlling dust collection and particle department through the scrubber system. Results show that the vibrating mesh conditions and higher water flow rates led to an improvement in the collection efficiency of the scrubber system. Compares to a static-mesh, the vibrating mesh FBS showed a significant reduction in pressure drop across the mesh screen indicating lower air loss through the test duration. Overall, the findings confirm that vibrating filter mesh reduces filter clogging while maintaining high collection efficiencies. This research supports further technological advancement in mine dust mitigation technologies.

Dec 27, 2023

By John Hillman

The reconstruction of the British mining industry which began shortly after nationalisation in 1947 continued until well into the 1950s with the support of the Government. This support diminished as plentiful supplies of cheap oil from the middle East made many coalmines un¬economic. The planned rundown which followed was foreshadowed in the 1965 White Paper which outlined the financial measures to speed the closure of uneconomic collieries. The decline appeared likely to continue into the 21st century as the role of coal in UK energy requirements became less and less significant. This policy was well on course in 1973 when the oil crisis led to a complete reappraisal of the position. A Government examination in 1974 of the coal industry then supported the National Coal Board's general strategy of major capital investment as set out in its 'Plan for Coal' (1). Plan for Coal aimed at halting the intended decline in the industry's capacity and restoring it to 120 million tons of deep mined output in 1985. The Plan proposed action along the following lines (a) increase output by major Investment in construction at existing collieries with sufficient reserves to give 13 million tons per annum extra output (b) invest in extending the life of existing collieries by gaining access to additional reserves, to give 12 million tons per annum extra output © by sinking new mines, produce an extra 20 million tons per annum. Figure 1 shows the contribution (2) expected from each of the three sources up to 1990/1. The measures taken were expected to offset the inevitable exhaustion of existing pits and provide the additional 40-42 million tons capacity needed to restore annual output in 1985 to 120 million tons. Parallel to this programme for expansion of deep mined output, opencast mining was to produce 10-15 million tons per annum. A recent review however, indicates that new mines are likely to yield only 10 million tons per annum by 1985, mainly owing to the need for public inquiries into proposals for new mining and the resultant fact that planning approval takes 'much longer than originally expected. Nevertheless the objective of 120 million tons deep mined output by 1985 is still considered achievable of means of a greater contribution from existing mines. This programme was originally estimated to cost 2,-56 million dollars at March 1974 prices, out revised estimates of the cost of Plan for Coal indicate that it may now exceed 6,960 million dollars at mid-1977 prices. Plan for Coal includes a substantial investment in new and reconstructed coal preparation plaints. At mid-1977 prices more than 322 million dollars Ls so be invested in coal preparation projects in the UK, at the rate of, about 70 million dollars a year. This rate 3f capital expenditure has been maintained over the past three years and covers investment in a wide range of coal preparation coals for the power station, coking and domestic markets.

Jan 1, 1978

By J. Kretschmann, R. Amiri

"The Sangan iron ore mines (SIOM) in Iran are located in a remote area near the Afghanistan border. The SIOM deposit contains a geological resource of 1.1 Gt (1.2 billion st) of mostly magnetite with an iron (Fe) grade ranging from 27 to 61 percent. The Iranian government plans a stepwise development of SIOM in cooperation with private companies to satisfy the growing demand of the Iranian steel industry. Sustainable development is an influential ethical concept in mining all around the world. However, a successful realization of a sustainable development plan is not possible without a foundation on the core values and regarding the specific circumstances of the individual mining country. This paper introduces SIOM and describes sustainable development activities undertaken during its development. From the early stages, these activities have been realized to secure the acceptance of the stakeholders in the region. To evaluate their effects, indicators were filtered to the ones with the highest priority for the stakeholders. Moreover, Islamic values and Iranian regulations and their influences on SIOM were analyzed.Mining industry in Iran The Islamic Republic of Iran holds 68 types of minerals, including salt, sand and gravel, gypsum, chrome, lead, zinc, copper, coal, gold, iron, etc., with 33.5 Gt (37 billion st) of proven reserves and more than 51.7 Gt (57 billion st) of potential resources. Iran is ranked among the 15 major mineral rich countries (1). In accordance with the statistics published by Statistical Center of Iran (SCI), there were 5,316 active mines and 84,922 employees in the whole country in 2012-2013 (2) and, according to the statistics published by Central Bank of Iran (CBI) in 2011, the share of this sector in the gross domestic product (GDP) of the country was 1.2 percent (3)."

Jun 1, 2016

By J. C. Haptonstall

Stoping methods at the Homestake Gold Mine through 1978 had been traditional overhand cut-and-fill, shrinkage or square set. High costs, low-productivity and safety concerns prompted the mine to search for and adopt a "bulk mining" technique as a primary stoping method. Blasthole stoping had been successful but it is very development intensive, and was not suitable for the mine's accelerated mining program. Vertical Crater Retreat stoping was selected, and today accounts for over 50% of ore production. Case histories representing both the successes and failures of VCR stoping in the mine are discussed.

Jan 1, 1986

By C. S. Compton, T. E. Marshall, D. R. Dolinar, R. Grau, L. J. Prosser, A. T. Iannacchione, D. C. Oyler

Unrecognized roof beam failures resulted in 69% of the falls of ground injuries occurring in underground U.S. stone mines from 1990 to 1996. Field investigations at 45 underground stone mines suggest that excessive beam sag or deflection is the principal unstable roof behavior characteristics associated with miner injuries. Laminated roof structures, common in most underground stone mines, allow the immediate roof to separate into distinct beams. The timely determination of roof beam deflection could provide stone miners with advanced knowledge of potentially hazardous roof. Existing methods to assess roof conditions consist of observational and monitoring techniques. Observational techniques include open drillholes, sounding, wedging, and borescopes. Monitoring techniques include scratch tools, extensometers, telltales, Miners Helpers, and Guardian Angels1. Miners have difficulty utilizing these techniques because roof beams are located an average of 8 m(26.5 ft) above the miner’s head in environments that are dark, dusty and noisy and these deflections are measured in millimeters of movement. To help mitigate the potential for falls of ground injuries to underground stone miners, the National Institute for Occupational Safety and Health (NIOSH) has developed the Roof Monitoring Safety System (RMSS) to aid in monitoring dangerous levels of roof beam deflection. If enough site-specific information is collected and placed on highly visible maps, individual mining operations could establish general guidelines for responding to specific roofbeam deflection occurrences. This would help in implementing a pro-active, comprehensive roof control plan for reducing falls of ground injuries.

Jan 1, 1999

By Ashutosh Kumar Bharati, Rajesh Rai, B. K. Shrivastva, Arunava Ray

Opencast mining method is one of the most productive mining methods available at present, which is generally associated with the removal and handling of a huge amount of overburden material. Hence it imposes a serious challenge of dump slope stability and that too, mainly in the case of internal dumps. The study is primarily focused on the stability of dragline dumps and the subsequent development of a rating system to classify the stability of dragline dump profiles. Initially, numerical simulation is executed through RS2 (Version 9.0) software, which is based on the finite element method. Thereafter, simulation results are used to perform the sensitivity analysis of various geometrical and strength parameters of the dump slope, and the two-factor interaction method is also incorporated into the analysis, this method included the dependency of closest parameters while calculating the sensitivity index, and thereafter, it is converted into the rating of each parameter. The strength parameter like the angle of friction and the geometrical parameter like slope angle of the bench between the coal-rib roof and dragline sitting level are assigned the highest rating value. Based on the overall rating of the dragline dump, the dump slopes are classified into various classes, viz., very safe, safe, vulnerable, unstable, and highly unstable. This classification proved to be handy in the quick estimation of the stability of the dragline dump and also gives an idea of designing optimum and safe dragline dump profiles. The developed classification chart has been validated by the previous research works and has found similar results.

Feb 15, 2021

By Perry M. Jones

An important operational and environ-mental concern associated with in situ leaching activity is the balance between solution injection and recovery. Solution balance is an important indicator of the capability of a particular in situ hydrologic design to efficiently and effectively control and distribute leach solution within the target ore zone. The Bureau of Mines is investigating the factors that affect solution balance in an unsaturated ore deposit, as part of ongoing in situ leaching hydrologic investigations at the Cyprus Casa Grande Mine. Flow conditions in 58 injection and recovery wells at the Casa Grande In Situ Leaching Test Facility were monitored daily over an 18 month period. Changes in the balance between injection and recovery occurring during this period were evaluated using analytical and statistical methods. The results permit estimation of solution distribution and the volume of solution stored in microfractures and interstitial pore spaces within the deposit following 18 months of leaching. Approximately half of the difference between well injection and well recovery can be accounted for by solution movement through macrofractures into nearby drifts. Estimates of solution storage within the ore deposit at the end of 18 months of leaching range from 3.5 to 6.9 pet. of injected solution. These results also suggest that 13 to 26 pet. of the fracture/interstitial pore space in the deposit was saturated with solution after 18 months of leaching.

Jan 1, 1991

By ANDREW HOBERT, Maoming Fan

Low-cost and environment-friendly green froth flotation modifiers of plant origin, such as guar gum, carboxymethyl cellulose (CMC), and starch, have been used in flotation of some sulfide and non-sulfide minerals. Certain green modifiers may be used as dispersants, flocculants, depressants, or activators. The functions of green modifiers in froth flotation highly depend on their molecular composition, flotation slurry chemistry, hydrodynamic conditions, etc. This paper will discuss the fundamentals, current applications, and potential new applications of guar gum, CMC, and corn starch in valuable minerals (phosphate, potash, and iron ores) and the unwanted gangue minerals (silica and clay minerals) flotation. In this study, improved flotation performance was achieved using some green modifiers in flotation columns with the more plug-flow environments than in conventional flotation cells approaching perfectly mixed conditions. This is because the hydrodynamic cavitation-generated ultrafine bubbles could help bridge ultrafine particles to form aggregates, promote the formation of flocs, and increase the stability of flocs.

Jan 6, 2022

By J. R. Goode

The first part of this paper comprises a general discussion on refractory gold ores and their mineralogy processing methods now in use are reviewed and test procedures for the important methods are discussed. Factors determining the selection of a process are outlined. The paper includes a survey of production plants using pressure oxidation, roasting and biological oxidation to process refractory ore. The status of processes under development is briefly reviewed.

Jan 1, 1993

Ground falls (roof and rib) have historically been responsible for nearly 50% of all fatalities in bituminous underground coal mines. In recent years, the number of some annual ground fall fatalities has approached zero, indicating that significant progress has been made. On the other hand, the twin disasters at Crandall Canyon in 2007, in which nine miners perished in violent coal bumps, provided a stark reminder that complacency is premature. One important success has been a great reduction in the number of miners killed inby roof supports. Throughout the 1990s, these accounted for nearly half of all roof fall fatalities, but there have just been two inby incidents since 2005. Progress has also been made in pillar recovery, where there has been just one fatal incident since 2005. On the other hand, more than 300 miners continue to be injured each year by rock falling from between supports and 100 more are injured by rib falls. Together, these two categories also account for a large percentage of recent ground fall fatalities available technologies such as roof screen, rib bolting and inside control roof bolters could reduce injury and fatality rates if they were used more widely. Further advances in these areas will likely be the next big advance in ground control safety.

Jan 1, 2011

By Howard S. Avery

The wear resistance of an alloy in heavy mill liner sections may be quite different from that of the same alloy in small balls and other light sections. This fact was established by means of a well standardized laboratory test that was validated for ball mill service. The salient properties of mill liner alloys, including austenitic manganese steels, medium alloy hardenable steels, and abrasion resistance cast irons are described, critically compared, and quantitatively ranked in terms of mill liner service. Where they are tough enough the high-chromium and other martensitic cast irons (e.g. Ni-Hard) promise maximum economy, but their section sensitivity is a limitation. To overcome this an extensive research program studied over a hundred experimental alloys, ranking them for heavy section abrasion resistance by means of the laboratory test. Computer analysis of the results provided a parabolic formula in terms of composition, the apexes of the parabolas providing optimum wear resistance. From the extensive family of parabolas one apex was selected for intensive study in experimental liners, pilot plant production, development of quality controls, and finally full scale production. This is a 19% chromium iron critically alloyed with nickel and molybdenum and hardened by heat treatment. It has excellent resistance to various types of abrasion, but its chief advantage is in heavy mill liner sections, where it has about 25% better abrasion resistance than the best of the quality controlled chilled Ni-Hards. It is also tougher. After production problems were solved, a number of ball and rod mill tests were arranged on a guaranteed performance basis. Two of these, representing about 138,000 pounds of mill liner castings, have recently confirmed in service the superiority that was predicted by the laboratory tests.

Jan 1, 1969

By Eduardo Pimenta de Melo, Rambabu Pothina, Rajive Ganguli

Sensors are everywhere in the mining industry, with sensor information being used to monitor and operate processes. Therefore, when sensor information is wrong, economic losses can occur. Unfortunately, sensor errors are usually not detected till they become large. This is problematic as most calibrate their sensors no more than once a year (Beamex, n.d.). Using principles of data mining, where all relevant information is tapped to glean hidden information, Pothina [8] designed an algorithm to detect errors in temperature sensors in gold stripping circuit in Pogo mine, Alaska. This paper continued his work by analyzing the behavior of the algorithm on baseline data and testing the algorithm in new data and under more rigorous conditions. It also made a change to the algorithm. The modified algorithm performed very well in the new data. It also worked well under the new error conditions. Three types of errors were seeded, a fixed additive error (+ 2%), a fixed subtractive error (− 2%), and a noisy, normally distributed error, with a mean value of + 2%. Artificially seeded errors were detected within about 20 gold stripping cycles. Inherent bias in operation impacted algorithm performance by increasing the number of cycles needed to detect errors. This paper reinforced Pothina’s [8] conclusion that when data mining approach is used, sensor errors can be detected even when they are pretty low. Significant economic losses can thus be minimized.

By F. M. Doyle, Jun Li, Milton E. Wadsworth, E. Peters

The Raman spectroscopic technique was used to study electrochernically oxidized pyrite in-situ. It was found that Raman-active sulfur oxidation products could be detected after a proper potential was applied to the pyrite electrode for a threshold period of time. There existed a minimum potential below which sulfur could not be observed. However, when the applied potential exceeded a certain limit, the intensity of the sulfur spectra decreased. The minimum potential at which sulfur could be observed changed with solution pH. For both acid (pH 2.7) and weak alkaline (pH 9.2) solutions, the minimum potential was 0.8 V/SCE, whereas for neutral (pH 6.5) solutions the value was 0.4 V/SCE. In strong alkaline solutions (pH 12.5) surface sulfur oxidation products could not be detected at any potential.

Jan 1, 1993

By K. M. Kim

An analysis was conducted to determine the effects of blasting on slope stability and downstream comminution processes. To investigate these effects, the extent of blast-induced rock damage and the magnitude of induced stress were examined. This analysis showed that damage may extend up to 500 m (1 MPa, tensile strength) into the rock mass, a distance that is sufficient to have an impact on the pitscale slope over the long term. In addition, the damage zone was calculated using the assumption that compressive waves are the primary agent in microfracturing during blast fragmentation. The calculated damage zone — up to 2.2 m (43 MPa, rock mass compressive strength) — indicates that microfracturing can occur throughout 20-50% of the entire rock mass in the blasting area, depending on the blast geometry, burden and spacing. Finally, an economic analysis was conducted by comparing two assumed blasts in a copper mine. The change in energy required for crushing and grinding was demonstrated using Bond’s law after increasing the blast energy from 180 kcal/t to 350 kcal/t. A total cost saving of 13.3% ($9.4M) was achieved with the increase. However, the benefits of blast-induced microfracturing must be balanced against the potential impacts to slope stability to ensure safe conditions, a top priority in mining operations.

Jan 1, 2014

By M. T. Ityokumbul

The use of yield index, the product of the yield and gangue rejection, for the optimization of separation processes is presented. The characteristics of the yield index are contrasted with separation efficiency calculations and applied to laboratory froth flotation and washability data. Data from continuous froth flotation tests conducted to "optimize" froth flotation using process optimization software were also evaluated with the yield index procedure.

Jan 1, 1997

By Andrea Brickey, JUDITH BUABA, RICHARD AMOAKO

The global mining industry has recorded significant declines in accident and injury rates attributed to the advancement in technology, increased enforcement, and safety consciousness. A goal of the mining industry is to achieve zero injury and occupational illness on all mine sites, prompting increased research into ways to further reduce mine accidents. A machine learning technique known as multiclass logistic regression is applied on a 10-year injury dataset from the Mine Safety and Health Administration (MSHA) to determine a miner’s susceptibility to a class of injury and to help identify significant risk factors associated with different classes of injury. The data is aggregated based on injury classification to provide statistically relevant results. The analysis identifies specific risk factors that influence a mine worker’s susceptibility to a given class of injury, i.e., non-fatal with no days lost or restricted activity, non-fatal with days lost and/or days of restricted work activity, and fatal and total permanent or partial permanent disability. These factors include miner’s age, mine type (coal vs. non-coal), experience on the current job (years), shift start time, employment type (operator vs. contractor), mining district, and type of accident. The results of the analysis indicate that a miner’s experience on the job, i.e., the number of years worked in a current job, is a significant risk to injury occurrence, even for those with decades of total mining experience. We further show the differences and similarities between the surface and underground mine incidents.

Nov 3, 2020

By K. E. Yassin, A. Yehia, M. Amar

Phosphate fines (size ≤ 45 μm), from the Red Sea Region, Egypt, were subjected to beneficiation process to recover the phosphorite grains from such fines. The sample is characterized by low P2O5 content (18.78%) associated with gangue minerals such as silica (22.77% SiO2) and carbonate (2.01% MgO) indicating its low grade. Phosphate pre-concentrate was prepared by gravity separation, using Falcon, where 63.37% of the phosphate, with 24.94% P2O5, 1.46% MgO, and 10.54% SiO2, was firstly recovered reducing the mass flow to the subsequent beneficiation process. However, the gravity tailing was the feed for the reverse phosphate flotation using bench scale column flotation where oleic acid was used as a carbonate gangue collector and amylase enzyme as a phosphate depressant. Under the appropriate flotation conditions (0.1% amylase, 5·10−4 mol/dm3 oleic acid, and temp. 30 °C), phospho-concentrate assaying 0.61% MgO, 13.14% SiO2, and 27.85% P2O5 with a P2O5 recovery of 55.45%, was finally obtained without the use of expensive depressants, e.g., phosphoric acid or sodium silicate. A tentative flow sheet for the whole process was postulated.

By Hans Ehdwall

INTRODUCTION Investigations of radon and radon daughter concentrations in Swedish [non-uranium] mines started in the late 1960's. The first screening measurements showed that the average annual exposure to radon and radon daughter products was 4.7 WLM. The main reason for high radon and radon daughter concentrations was inefficient ventilation and radonrich water entering the mine. In the radon regulations worked out later it was stated that no miner should be exposed to more than 60 000 pCi h/1 equilibrium equivalent concentration of radon annual exposure, corresponding to 3.6 WLM. Now, 1981 the situation has changed considerably. From the average annual exposure of 4.7 WLM in 1970 it is now only 0.7 WLM. Sweden has up to now had only one [uranium] mine and the work there has only been investigative. However, there are plans for a commercial uranium mine in another part of Sweden. The radon problems in these mines are widely different depending on the mineralogy. NON-URANIUM MINES The radiation problems in Swedish mines were not recognised until the late 60's. The first radon and radon daughter measurements were made in some sulphide ore mines in 1967 (1). The radon and radon daughter concentrations were surprisingly high for non-uranium mines. In order to have a complete picture of the radon situation in Swedish mines the National Institute of Radiation Protection (NIRP) decided to make measurements in all, at that time about 60 mines (2). To get results as fast as possible measurements on radon gas seemed most appropriate to start with. Sampling was made by mailing a number of evacuated 4.8 litre conventional propane containers from NIRP to each mine. The containers were then opened at the place of interest. After sampling the containers were sealed and then mailed back to the institute for measurement. The measurements were made in ionization chambers. This method only gave the radon concentration and the radon daughter concentration was estimated by multiplying the radon concentration by an assumed equilibrium factor. The equilibrium factor is defined as the ratio of the total potential alpha energy for the given daughter concentration to the total potential alpha energy of the daughters if they are in equilibrium with the given radon concentration. The results of this first preliminary survey indicated that a great many of the Swedish miners probably had an annual radon daughter exposure of more than 3.6 WLM. As the radiation exposure in non-uranium mines was not regulated in either the Swedish Radiation Protection Act or the Swedish Labour Protection Act work was started on special radon regulations. A lung cancer mortality study was also started. To check the results of the first survey and to get experience and knowledge of radon problems in mines, it was decided that personnel from the NIRP should visit each mine for a detailed investigation of radon and radon daughter concentrations starting with the ones with the highest radon concentrations. The main reasons for these so-called "basic measurements" were: 1. To estimate the doses received by Swedish miners 2. To find the sources of the high radon and radon daughter concentrations 3. To find appropriate counter-measures 4. To determine the most typical equilibrium factor for each mine. Unlike most uranium mines the reason for high radon concentrations in non-uranium mines is seldom the occurrence of highly radioactive minerals. The main sources were found to be waste-rock and radon-rich water. In order to filter and warm up the inlet air, especially in winter time, it was very common at that time to suck the air through broken wasterock. By doing so the air was contaminated with radon from the waste-rock and radon-rich water in it. It is noteworthy that the radium and uranium concentration in the waste-rock is relatively low. The uranium concentration is only of the order of 15 - 20 ppm. The action to prevent this contamination of the inlet air was to change the direction of the ventilation and in the case of radon-rich water entering the mine the action was to prevent the air coming into contact with the water. The first calculation of the radon daughter exposure of Swedish miners was based on radon gas measurements. The radon daughter concentration was estimated by using an assumed equilibrium factor of 0.5. Later when the mines were visited by institute staff it was possible to compare the assumed equilibrium factor with the measured ones. It was found that the factor varied from 0.15 at the air inlet to 1.0 at the air outlet and the average equilibrium factor on workplaces for almost all mines was between 0.4 and 0.6. The result of the exposure calculation in 1970 showed that more than 40 % of the miners had an annual radon daughter exposure of more than 3.6 WLM. The overall average was 4.7 WLM and the maximum annual expo-

Jan 1, 1981