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By Andre Vien, SANDRO DAVID DI, Frank Laroche

In 1994, Claude Bazin presented a novel method for linking size data from milling models to the feed input of flotation models. It was achieved with the simple addition of a metal/mineral distribution–size distribution relationship. This allows grade and recovery predictions from a varying mine plan through product size distribution predictions. 

By Kenneth Deaton

"A packaged cooling system is being developed for refuge alternatives (RAs) that will meet U.S. Mine Safety and Health Administration (MSHA) qualification guidelines for deployment in coal mines. This system will allow full occupancy ratings of deployed refuge alternatives to be maintained by keeping the apparent temperature inside the RA from exceeding 35° C (95° F).Research investigations sponsored by the National Institute for Occupational Safety and Health (NIOSH) have shown that heat buildup caused by occupants in mobile RAs will result in apparent temperatures that exceed the 35° C (95°F) limit established in 30 CFR 7.504.Figure 1 shows the temperature rise in an inflatable RA having an occupancy rating of 23 people at an 27° C (81°F) mine air temperature. This plot shows the need for either heat mitigation or occupancy derating in warmer mines.The Refuge Alternative Environmental Control System (RAECS) will be designed to provide enough heat mitigation to eliminate occupancy derating due to heat buildup over the 96-hour refuge period.This article describes testing of a prototype cooling system that uses proven vapor compression technology powered by batteries to keep the RA interior below the 35° C (95° F) limit for a minimum of 96 hours. Testing of this prototype cooling system was conducted in December of 2015 and January of 2016 at the NIOSH test mine in Bruceton, PA. December testing was conducted at the prevailing mine ambient temperature — approximately 15.5° C (60° F). The January testing was done with the mine ambient temperature in the immediate vicinity of the RA artificially elevated to 29° C (85° F). Realistic heat gain from RA occupants was provided by human simulators developed by NIOSH. These simulators add sensible heat corresponding to human metabolic processes and latent heat generated by respiration and perspiration. Heat generated by CO2 scrubbers was also added. Results of cooling system performance using these simulated loads at the two ambient conditions are described. A comparison of the apparent temperature in the RA using two different control techniques is also presented. This work documents the ability of the cooling system to meet the established apparent temperature limit at the full occupancy rating of a six person RA. A new cooling system intended for use with existing RA deployments and having greater cooling capacity is currently in development and will be described.Expendable desiccant modules were also used to provide supplemental water vapor removal and their role in mitigating cooling load is addressed."

Dec 1, 2016

By Bo Kyung Sim, Dae Young Kim, Jae Won Lee, Ebrahim Farrokh

"The contract C931 of the downtown line 3 (DTL3) mass rapid transit system of Singapore includes half a kilometer twin tunnels running between Mattar and Macpherson stations. An EPB TBM of 6.6 m diameter was used to excavate both of these tunnels. The tunnels pass predominantly through a very stiff silty/clayey fine to coarse sand from the Old Alluvium Formation. In this paper, TBM selection procedure, TBM performance evaluation, and some of the tunneling operation challenges with the TBM in the Old Alluvium Formation are discussed. In this regard, soil conditioning trial tests were performed to resolve the clogging issue and low TBM performance. Thorough TBM operational data analysis was performed to detect high soil abrasivity, backfill grout leakage, and foam nozzle clogging. The results of some of the applied modifications are also discussed to show their effects on the TBM performance improvement. Finally, some of the challenging TBM design specifications are also discussed. INTRODUCTION Tunneling and underground construction in Singapore has been increased in the last 30 years due to the demand for public transportation systems and the limited space above the ground for the infrastructure development. Downtown Line Stage 3 (DTL3) Mass Rapid Transit (MRT) system has 21 km total route length and consists of 16 stations. DTL3 Contract C931 runs between Mattar station and MacPherson station. The construction works in this contract include twin bored tunnels of 6.6 m diameter and 450 m length and a TBM launch shaft of nearly 30 m depth. The tunnels run in urban area below Kallang Paya Lebar Expressway tunnel (KPE), Kallang River Canal, and Circuit Link. Figure 1 shows the alignment of the tunnels in between two subway stations. The tunnels were excavated by an earth pressure balance (EPB) tunnel boring machine (TBM) through the Old Alluvium (OA) formation, a challenging geological formation in the sense of its abrasivity and stiffness. The Old Alluvium Formation mostly includes stiff to very stiff clayey and silty sand. The soil is un-weathered cemented to semi-cemented dense, mainly with a low permeability coefficient of 10-7 m/s and high SPT values (almost 100). The soil is classified as “very high abrasive” according to NTNU SAT test classification (SAT value of 15 to 42). The construction of the first tunnel drive was slowed down due to a number of factors related to soil conditioning and TBM operation. In this regard, a thorough analysis of the TBM operational parameters from the excavated sections was performed to identify the problems in time to prevent unwanted construction delays in the subsequent sections. In the following sections, a brief overview of the criteria used for TBM selection is presented at first. Then, the results from the analyses and tests performed to assess soil conditioning, grout leakage, and soil abrasion are presented. The last section discusses some of the tunneling design features that challenged the TBM operation in this project."

Jan 1, 2016

By R. L. Murray-Smith, M. E. A. Robertson, M. J. Prosser

The AZTEC is an automated instrument, for the non-destructive analysis of gold ore samples, which uses energy dispersive X-ray fluorescence (XRF), This is achieved by X­ray tube excitation of gold K X-rays. The first production instrument was commissioned at Vaal Reefs Chemical Laboratory (VRCL) in November 1987, followed by two second generation instruments in 1989. AZTECs have also been installed at Western Deep Levels and at Anglo American Research Laboratories (AARL). A further unit is being assembled at Ashanti mine, Ghana. At VRCL the AZTECs are used primarily for the evaluation of mine (stope) samples. Each of the production instruments can routinely analyse 650-700 samples per day, with the aid of a robotic sample transfer system. Various techniques have been evaluated for increasing sample throughput. For example, higher value samples can be counted for a shorter time to obtain the necessary relative precision.

Jan 1, 1991

By Ralph S. Mattson, Frederick T. Hancock

GENERAL DESCRIPTION The Granduc mine is situated on the west side of Granduc Mountain in the Leduc River area of north¬west British Columbia about 51 km (32 miles) north¬east of the town of Stewart. Entry is by means of a 16.6-km (10.3-mile) long haulage adit from the concentrator site at Tide Lake. The principal ore mineral is chalcopyrite. Ore mineralization varies from dis¬seminated to irregular semimassive stringers generally associated with highly folded and sheared biotite-rich sections of metasediments. The ore bodies vary in width from 3.6 m (12 ft) to over 18 m (60 ft) in a zone some 800 m (2600 ft) in length dipping at an average of 75° to the west and plunging steeply to the south. Capping varies in thickness from 0 to over 300 m (1000 ft). SELECTION OF MINING METHOD In the early stages of planning the primary require¬ments to be met by the mining method were: (1) suit¬ability to high output and productivity, (2) adaptability to mechanization in order to minimize operating costs, and (3) minimization of the costs and the development period for preproduction. Considering these, the selection of a mining method was narrowed to a choice be¬tween sublevel long-hole stoping or sublevel caving. A method utilizing backfill was ruled out as uneconomical, especially since mill tailings would not be readily available. At the time, the following facts about the ore body were recognized: (1) the average width of the ore body was 12 m (40 ft), (2) 30% of the stoping blocks were in areas of 6 m (20 ft) or less in width; (3) in many cases, ore zones were located in parallel lenses or were irregular in shape; (4) major faults were present in the hanging wall close to the ore, with minor faults within the different ore zones; and (5) the ore itself was lami¬nated with fractures and jointing crossing such bedding planes. Considering these factors, the two stoping meth¬ods were compared. Sublevel Open Stoping The characteristics of sublevel open stoping are: 1) It is generally not used for widths below 6 m (20 ft) since the ratio of waste development increases as level intervals are reduced to insure good control of long-hole drilling. 2) This method is used where ground is relatively competent. Otherwise more pillar support is required, increasing ore losses or mining costs with pillar recovery. 3) The method is not easily adaptable to parallel ore zones separated by relatively narrow waste bands. 4) Close control of pillar and stope design will per¬mit some degree of dilution control. Sublevel Caving Sublevel caving is characterized by the following: 1) The method can be used for a variety of ore widths, recognizing that 3 m (10 ft) in width may be an economical minimum. 2) It is particularly favorable where wall rock is weak. 3) The method has flexibility so that irregular ore bodies or parallel lenses can be mined. 4) The need to leave pillars for support is eliminated 5) Dilution and tonnage recovery are interrelated. With concern about the stability of the hanging wall and the geometry of the ore bodies, there was a growing favor to choose the sublevel caving method. However, this method was relatively new to Canadian mining practice. Teams were sent to Sweden to tour mines using sublevel caving and to study their methods and any extraction and dilution problems. The objective was to develop cost and performance data and ore recovery grades and tonnages for feasibility comparisons. Next, detailed mine plans were developed covering a three-year period using sublevel caving and sublevel open stoping. Extensive analyses of the development and production costs for each method were made. With sublevel caving, ore losses of 10% with a 20% dilution rate were used. The same ore loss was estimated for sublevel open stoping with dilution varying from 11 to 15%. As a result of these studies, the decision was made to use sublevel caving partly because slightly lower op¬erating costs, together with higher productivity, offset the adverse effects of higher dilution; and partly be¬cause, at the start, there would be less risk in being able to handle the ground, and the method offered more flexi¬bility. Open stoping was not ruled out for testing at a later date. It was recognized that some other method of mining would be required for mining the ore extensions below the glacier, where surface subsidence must be avoided. During the operating life of the mine, open stoping was tested in narrow width ore zones and found un¬satisfactory due to wall failure. A cut-and-fill stope, utilizing run-of-mine waste from development, was also operated on a trial basis. A sequence was finally estab¬lished that permitted ground control; however, produc¬tivity needed considerable improvement in order for open stoping to be considered as a sole alternative for mining below the main haulage level. At the termination of operations in June 1978 over 13 000 000 t (14,500,000 st) had been extracted from above the main haulage level at production rates varying from 3270 to 7250 t/d (4000 to 8000 stpd). Sublevel Caving Methods Two basic methods were used, transverse and longi¬tudinal. The transverse method was used mainly in the upper C ore body in widths greater than 18 m (60 ft), and occasionally as an alternative to the longitudinal method in localized areas where strike faulting inter¬fered with normal development. The longitudinal method was used in ore bodies varying in width from 3.6 to 18 m (12 to 60 ft), with a multilongitudinal varia

Jan 1, 1982

By Craig J. Nelsen, J. Leroy Schutz

The Carache Canyon Prospect consists of a hypabyssal breccia pipe which intrudes a gently dipping sequence of Cretaceous Mancos Formation and Mesa Verde Group shales and sandstones. This sedimentary sequence is intruded by a number of Oligocene (34-30 Ma) calc- alkaline sills. The formation of the breccia pipe and its subsequent gold mineralization is related to the emplacement of an alkaline intrusive complex dated between 30-26 Ma. An exceptionally large block within the pipe provides compelling evidence that brecciated sills and sedimentary rocks form a relict stratigraphy that collapsed 400-800 ft (120-240 m) below their counterparts outside the pipe without substantial changes in stratigraphic position. Native gold and sulfide mineralization within and around the pipe occurs in fractures that are primarily controlled by competency differences between sills and sedimentary rocks and secondarily controlled by ductility contrasts between the relatively brittle Mancos Formation hornfels and the more ductile Menefee Group shales. The resulting pattern of gold mineralization is a series of stacked, tabular, mineralized sills and sandstone separated by barren shale in an annular region along the southwest margin of the pipe. A series of near-vertical, subparallel, sill and sandstone breccias separated by barren or weakly mineralized shale breccia occur within shallow levels of the pipe.

Jan 1, 1990

By Nathaniel Wollman

About nine years ago the President's Materials Policy Co commission under the chairmanship of William Paley released its report entitled "Resources for Freedom." Not long ago the Senate Select Committee on National Water Resources under the chairmanship of Senator Kerr released a study entitled Water Supply and Demand. Both reports contained projections of future water user the report for 1975, and the Select Committee print for 1980 and 2000, It is my task to compare the two reports. In the initial letter from Mr. Kazmann arranging for thus session he suggested that I present "a critical evaluation of the Paley Report's section on water ?ten years later,? which might include a comparison of predicted and actual water-use rates and would spotlight the assumptions which lead to errors in the predictions , . ." It is possible to compare the results of both studies, and it possible to explain art of the differences that can be found between them. The remainder must remain shrouded in doubt, since the Paley Report did not give all details of the methods used in estimating the com¬ponents of water use.

Jan 1, 1961

By J. A. Herbst, H. R. Cooper, R. J. Brison, R. W. Harper, K. P. Ananthapadmanabhan, W. C. Hellyer, B. H. Bergstorm, D. M. Hopstock, H. Rush Spedden, W. C. Hellyner

Scope of Section The physical characterization of an ore or mineral mixture in such detail that mineral processing may be effectively applied to create more useful products is one part of the technology of sampling and testing. In addition, the evaluation of the multitude of separation processes, both physical and chemical, combined with ancillary opera¬tions and performed on samples at a laboratory scale, is another part of the testing procedure. Although analytical techniques provid¬ing quantitative determinations of elemental compositions are an im¬portant and necessary supplement to any testing program, that field of endeavor is a well-developed, separate entity beyond the scope of this Section. Sampling procedures cover the practice of selecting representative quantities of test materials in the field to evaluate mineral deposits, and extend to techniques of obtaining process samples in slurry and bulk ore as part of process evaluation. The same techniques are appli¬cable, of course, to standardized methods of metallurgical accounting. Sampling is also an important facet of successful on-line measurement of slurry properties such as assays and particle size in the practice of automation. However, the sampling methods in this case can be different from those needed for precise testing and accounting pur¬poses. Sampling was well defined in Taggart's Handbook as "the opera¬tion of removing a part convenient in size for testing, from a whole which is of much greater bulk, in such a way that the proportion and distribution of the quality to be tested (e.g., specific gravity, metal content, recoverability) are the same in both the whole and the part removed (SAMPLE). The conditions of the more stringent definition, that the sample shall be completely representative of the whole as regards all aspects save bulk, are practically never fulfilled when heterogenous mineral mixtures are sampled." Taggart's Handbook of Mineral Dressing, Section 19, "Sampling and Testing," continues to be a very useful compendium of principles and techniques, particularly as applied to manual methods for han¬dling samples. In the intervening years since its last revision, substan¬tial progress has been made in both the theory and the techniques; thus, this current contribution is designed to be an extension of the earlier work, yet complete enough to provide procedural guidelines to methods of continuing validity. Although basic principles of testing remain unchanged, the meth¬ods of applying those principles have undergone great change. New testing equipment has become available, yielding a greater precision of results. Testing methods have been refined which, in many cases, allows for a valid scale up of flowsheet designs without the necessity of costly pilot or prototype plant development. Justification for a Sampling and Testing Program* The proper evaluation of a mineral property is a complicated process. It often is done on a team basis with the exploration geologist, mining engineer, mineral process engineer (extractive metallurgist), economist, and product utilizer (ceramist, chemical engineer, chemist, metallurgist, etc.) all represented. The mineral processing engineer is a key member because he can provide a go no-go answer to the question, "Can this ore be processed economically?" In addition, his judgment on the response of even the initial sample to processing can be an invaluable guide to the proposed process and its mutations. Since the capital investment and operating costs are so sensitive to the type of process used, the judgment of the mineral processing engineer forms an important part of the early evaluation of the profitability of a new mining venture. This early estimation of the processing approach is all the more important since, for some metalliferous ores, the capital investment for mill, smelter, water supply, and tailings disposal exceeds the cost of the mine. At present, for commodities such as coal, the reverse is true. There is a smaller capital investment for the coal preparation plant than for the mine. In general, both the capital and operating costs escalate very rapidly in the following order: gravity concentration of alluvial deposits (e.g., dredging), gravity concentration of lode ores, flotation, simple hydrometallurgical processes, and complicated hydrometallurgical and pyrometallurgical processes. The latter two processes may cost 100 to 1,000 times more per ton treated than processing of a simple alluvial deposit. This dictates that the valuable mineral be concentrated as much as possible to minimize the total tonnage treated in the more expensive operations. Within each of these ore-processing categories there is a wide range of process sophistication and hence of capital and operating costs. Perhaps the greatest variations are in the chemical processing of ores. These range from the simple leaching of a copper waste dump to very sophisticated hydrometallurgical processing of ores in increasing order of cost. Because the overall cost of a mining venture is so sensitive to the process selected, it is essential that the mineral¬ processing engineer evaluate the ore in the initial stages of the development of an ore body. As better samples become available, he must continually revise his estimate of the process to be employed. The sampling and testing process must be a necessary and continuing activity throughout the life of a project. One fundamental aspect of nearly all mineral deposits is the inherent variability of the mineral assemblage and composition. A utopian goal of any processing plant is to have the process so flexible and so well-defined that an automatic process control system can optimize the economic recovery at all times. Since this is rarely possible, a continuing or at least a periodic testing program is desirable. Variables to be considered are: 1) Ore body inconsistencies. 2) Zonal variations. 3) Feed variations from mining plan or methods. 4) Variable performance of processing units due to wear. 5) Seasonal or other changes affecting processing conditions such as water quality, temperature, etc. 6) Changing market conditions for product quality or byproducts. 7) Technological advances in processing methods or equipment. 8) Increasing demands for environmental control. 9) Changing economics of competitive processes, of supplies, of capital investment vs. labor costs, etc. Sampling Procedures Preliminary Guidelines for Sampling Ore Deposit Studies. t At a very early stage, it is essential that the mineral processing engineer receive a sample of the ore. This may be a portion of the diamond drill core or a split from the trench, pit, churn drill, channel, or sheet sample. Even a grab sample may have some value, but it should be selected by the mining engineer or exploration geologist to be as representative of the ore body as possible within his best judgment. The sample delivered to the process engineer should have been subjected to as little prior crushing as is practical since fine sizes are difficult to concentrate. In addition, finely ground ores are readily oxidized and thus may give misleading test results. Furthermore, ground samples may make it impossible to eval¬uate either the comminution or the coarse-gravity concentration pro¬cesses. For an initial evaluation, a good deal may often be learned from a sample of only a few pounds if that is all that is available. To obtain the maximum information from a sample, its selection should be carefully discussed with the mineral processing engineer. In addi¬tion to being as representative as possible, the mineralogically distinct samples should be carefully segregated rather than composited. Since drastically different processes are often needed to process samples of differing mineralogy, considerable care and time must be exercised during this sample collection step. The segregation of samples showing gross mineralogical differences, such as the oxide caping on a sulfide deposit or a laterized zone, is easy. It is the more subtle changes in mineralogy that can often elude the field man. Yet these differences are often sufficient to dictate that an entirely different type of process be used. Obviously, the selection of these initial samples for ore testing and mineralogical studies requires that it be done by a man of consider¬able experience and judgment. While the mining engineer or field geologist is collecting these samples, he should also note and convey to the processing engineer such details as water quality and availabil¬ity, terrain, potential mill location, tailings-disposal areas, case of transportation, etc. As the property is developed with additional drill holes, adits, etc., larger and more representative samples will be required to develop the ore-processing flowsheet more adequately. If pilot-plant studies are required, a very large sample may be necessary. Though the type and amount of sample collected for flowsheet development investigations vary widely, gathering a representative sample of the ore body is a prime requisite for any process development effort. The process developed for any given ore body is only as reliable as the sample on which the flowsheet was based¬

Jan 1, 1985

By J. R. Dyni

The lacustrine Green River Formation of Eocene age in the Green River Basin, southwest Wyoming, contains the world's large resource of natural sodium carbonate in bedded trona (Na2CO3? NaHCO3? 2H2O). In this study, 126 billion tons are estimated to be present in 22 beds, ranging from 4 to 36 feet in thickness. These 22 beds underlie areas of about 50 to more than 775 square miles at depths ranging from about 650 to more than 3,000 feet below the surface. The total resource of trona ore in the basin for which drilling information is available is estimated to be about 149 billion short tons. The undeveloped deeper trona beds in the southern part of the basin may be exploitable by solution mining. Additional unevaluated sodium carbonate resources are present in disseminated shortite in strata interbedded with the trona and in shallow sodium carbonate brines in the northeast part of the basin.

Jan 1, 1995

By David R. Klug, Chris Gause, Troy L. Page

The Inter-Island Tunnel is one of two undersea tunnels which are part of the Boston Harbor Clean-Up Program. The tunnel is 7.7 km (25,160 ft) long, bored to a 4.3 m (14 ft) diameter in the Cambridge Argillite Formation and was finished to a 3.5 m (11.5 ft) diameter final concrete lining. The owner is the Massachusetts Water Resources Authority (MWRA) who is a regional authority responsible for overall design and construction. The tunnel's primary purpose is the transport of untreated effluent from Nut Island facility in Quincy, MA to the new treatment plant at Deer Island. Two sludge lines are installed for the transport of sludge from Deer Island to Nut Island. During construction the contractor encountered fault zones, fractured strata, and soft ground tunnel sections which liberated more ground water than was originally anticipated by the project designer. Ground water was controlled during mining and the placing of the final lining. The final lining of cast-in-place concrete had to be backfill grouted to cut off water inflows and provide contact for equal tunnel loading which is critical in long-term design. This paper deals with an engineered backfill grouting system which was the first of its type to be used in North America. It discusses the geological conditions encountered, the design criteria used in the evaluation process, and an indepth review of the various system components. It demonstrates how contractors can use knowledgeable and professional industry suppliers to solve difficult construction problems.

Jan 1, 1999

By A. Bruynesteyn

Many papers 'have been publihed on the role of the bacterium.Thiobacillus ferrooxidans in the oxidation of reduced sulfur compounds, and years of research have established the following environmental factors, which influence biological oxidation. 1. Oxygen is required for the oxidation of reduced sulfur compounds to the sulfate form. Sufficient oxygen. is available from the air, consequently, it is of importance to get the air to the bacteria. 2. The bacteria are active in the pH range 1.4 3.5. Maximum activity occurs at pH 2..0 - 2.8. 3. -In order to allow the bacterial populations to expand, certain essential nutrients must be available. Nitrogen has to be supplied for the production of body protein, and ammonia is the source most readily accepted by the bacteria. Carbon dioxide is the only utilizable source of carbon and the small amounts available in air are the primary source of this element. Small amounts of phosphate, manganese,, potassium,. calcium. and magnesium are also required. These trace minerals are available from the water and host rock in sufficient quantities. 4. Water is required as a carrier for the oxygen carbon dioxide and other nutrients, as-well as a solvent for the metallic sulfates produced. The bacteria are motile and the water allows the bacterial populations to spread:: bacterial activity ceases if the equilibrium relative humidity drops below 100%. 5.-optimum bacterial activity occurs at temperatures of 34 - 35°.C (93 -.95° F). All activity ceases above 45° C. At lower temperatures, activity its dramatically decreased, but there is still a significant amount at 10 - 11° C.- Thiobacillus ferruoxidans has been isolated from a mine with a year round temperature of 3° C. 6. Unpublished results from our laboratory show that biological leach ore oxidation takes place during the rest periods at the same rate as during the wet leaching periods providing there is a film of moisture remaining around the particles. Optimum conditions can be readily achieved in the laboratory, but application of our laboratory knowledge to commercial leaching dumps is riot as simple. Attempts to achieve improved environmental conditions for the bacteria in leaching dumps would require substantial capital investment and careful planning of leaching operations. This paper

Jan 1, 1972

By W. R. Reed, S. Klima, J. Driscoll, T. W. Beck

"Tests were conducted to determine properties of 4 foam agents for their potential use in longwall mining dust control. Foam has been tried in underground mining in the past for dust control and is currently being reconsidered for use in underground coal longwall operations in order to help those operations comply with MSHA’s lower coal mine respirable dust standard (1.5 mg/m3). Foams were generated using two different methods. One method used compressed air and water pressure to generate foam, while the other method used a low pressure air generated by a blower and water pressure using a NIOSH developed foam generator. Foam property tests comprising of a bottle shake test, foam expansion ratio, and water drainage were used to classify foams. The bottle shake tests determined that the lowest concentration of foam agent to be used was approximately 1.5%. Compressed air generated foams tended to have low expansion ratios (10-19) with high water drainage. Blower air generated foams had higher foam expansion ratios (30-60) with lower water drainage. Foams produced within these ranges of expansion ratios are stable and potentially suitable for dust control. The different foam agents produce foams with different material properties. Results of testing eliminated two foam agents for future testing because they had poor expansion ratios. The remaining two foam agents, while having different foam properties, seem to have properties adequate for dust control. These material property tests can be used to classify foams for their potential use in longwall mining dust control. INTRODUCTION About one-half of U.S. underground coal is produced by longwall mining. This technique allows for mining high volumes of coal, producing significant amounts of coal mine dust. This can lead to the overexposures to longwall miners and possibly occupational respiratory diseases black lung and silicosis, which have no cure and can be disabling or fatal. The only method to avoid these occupational illnesses is through elimination of exposure to respirable coal mine dust and crystalline silica (quartz). The current occupational exposure limit for respirable coal mine dust is 1.5 mg/m3 during each shift that a miner is exposed in the active workings of the mine or in mine facilities [1]. When respirable quartz is present, the mine must maintain an average concentration at or below 0.1 mg/m3. If the mine exceeds the 0.1 mg/m3 respirable quartz dust concentration, then the applicable respirable dust standard is reduced, calculated as 10 divided by the percent quartz present [2]. Respirable dust samples collected by U.S. Mine Safety and Health Administration inspectors for the five-year period from 2010 to 2014 found that 5.6% (53 of 943 respirable coal mine dust samples) of the longwall operators on the tailgate side exceeded the coal mine dust standard, and 4.3% (76 of 1768 samples) of the jacksetters exceeded the standard. If silica is present, the reduced standard is implemented, meaning that 11.4% (36 of 316 respirable coal mine samples analyzed for dust) of the jacksetters exceeded the reduced standard. Only 3.8% (18 of 478 samples) of the longwall operators on the tailgate side exceeded the reduced standard [3]. Had the 1.5 mg/m3 standard been in effect during this period, non-compliant samples would have been much higher. For example, over 15% of the samples collected at the tailgate shearer operator occupation exceeded the lower standard."

Jan 1, 2017

By Barbery G., A. Maachar, M. Bourassa

Laboratory testing for flotation circuit design is carried out on a routine basis in a large number of lab- oratories throughout the world. Usually a sample of ore, assumed to be characteristic of the material to be processed in a future plant, having a mass of the order of a few kilograms, is ground in a batch ball or rod mill, in the presence of water and some re- agents (for pH control, etc. .. ); after dilution to the adequate solid concentration, the resulting slurry is transferred into a laboratory batch flotation machine, having a volume which can vary from 0.5 to 8 10 -3m3. Various types of machines are in common use, manufactured by the same companies that market industrial size flotation machines. The operation of flotation in the laboratory is carried out by an experienced technician, who uses parameters such as impeller rotation speed, air flowrate addition, froth height, froth removal rate, in conjonction with reagent addition sequence, to "test" the ore. After flotation, the froth products ("concentrates"), and the solids remaining in the cell are filtered, dried, and analyzed for the various elements of interest. For flotation circuit design, a kinetic test is usually carried out, in which the froth products obtained at various times are collected to assess the time response of the ore in the laboratory machine with the combination of reagents used. Although this description is rather sketchy, it enables to outline some of the major difficulties in using laboratory data for flotation circuit design: 1. ore sample representatively 2. feed preparation procedure (laboratory compared to industrial scale) 3. laboratory procedures (type of ma- chine, type of operation, experimental errors. . . ) 4. flotation machine scale-up 5. flotation kinetic scale-up 6. flotation circuit design using laboratory data These various points will be discussed in the present paper, in order to stress the problem areas as well as possible solutions. ORE SAMPLE REPRESENTATIVITY It has been stated many times that one of the major problem in designing a flotation circuit is the representatively of the sample to be taken. Sampling of an ore body during prospection and evaluation is a specialized field, which has been progress in with the introduction of geostatistics. The application

Jan 1, 1986

By Someswar Kesh, Gavin J. Faulkner

Introduction Powered supports in longwall mines essentially operate under two conditions: 1. Static (non-working) 2. Dynamic (working) There is a variety of static conditions which may exist in a mine. These may include a temporary stoppage of work for a few minutes or stoppage of work for days because of strikes or lack of demand for coal. Studies on pressure behavior were made during brief stoppages of work (few hours) as well as a long stoppage of work (two weeks). In dynamic conditions, the pressure variation on supports is expected to follow a certain pattern. However, in many cases that does not happen. Pressure-time characteristics of each leg of the supports were studied. This paper discusses the research findings of the studies made on powered support pressure behavior under both static and dynamic conditions. Studies on Pressure Variation in Static Conditions It has been noted (Brenner et a1.,1975) that the pressure distribution on mechanized longwall faces is random in nature and changes with time and position along the face. Figure 1 and Figure 2 show the distribution of support pressures for the front left leg of supports in two different mines. In both the cases it can be seen that the variation of the support pressures along the face is indeed absolutely random. However, the nature of randomness in both the cases is very different. The data in Figure 1 (Mine 1) is much more random than the data in Figure 2 (Mine 2). This can also be seen from the standard deviation of the data (81 26 Pa versus 4000 pa) . By comparing the results of ten other sets of data the same conclusions can be arrived at. Since it was found that in Mine 1, the pressure on support legs was extremely random, batches of ten supports each were taken and the mean pressure plotted. This mean pressure line is shown in Figure 1. It was expected that this would smooth out the variation due to individual props and reveal a better picture of the support pressure distribution. For the case cited, the standard deviation is lowered to 5043 Pa from 8126 Pa which indicates that, on an average, the supports tend to take loads more uniformly, however for individual support legs the load on the supports is very random. The significance of this is that, if the supports are selected on the basis of theoretical load calculations alone (based on the roof block concept or some other concept), only the mean value of the support load required will be known. In the case of a pressure distribution with a high standard deviation, some supports selected on the basis of mean value will fail, while others will be supporting low loads. Hence, for proper selection of support capacity, it is not only necessary to know the average load but also the expected nature of the distribution. The only reason why the supports are supporting the roof in Mine 1 well is the fact that the yield load is above the highest level of load attained. If Mine 2, which has a much higher average load, had the same sort of standard deviation, the loads on the individual legs could easily exceed the support load. For example, if the pressure distribution followed a normal curve and the mean pressure was 36,700 Pa with a standard deviation of 9000 Pa, it is possible that 67% of the supports will be within the pressure range of 36,700 plus or minus 9,000 Pa, or 22,700 and 45,700 Pa. Many of the supports could then exceed the yield load of 40,000 Pa. It is clear then that the mean load is an inadequate criterion for estimating support resistance. The nature of the problem is statistical and such an analysis should therefore be made as far as pressure data are concerned. Statistical Analysis of Pressure Data. In general, in longwall mining and other geo-mechanics studies, it is extremely difficult to specify the exact nature of the distribution of a variable. Under these circumstances, probability

Jan 1, 1986

By David Caiden, Petros Fortsakis, Alan Hodgkinson

"1. INTRODUCTIONDo contractual practices affect project outcome? Can better contractual practices improve the likelihood of project success? The consensus amongst experts is that better contractual practices really do improve project outcomes. With this in mind the members of the International Tunnelling Association (ITA) Working Group 3 on Contractual Practices (WG3) pooled their collective knowledge to publish in 2011 the document entitled The ITA Contractual Framework Checklist for Subsurface Construction Contracts (ITA Working Group 3, 2011). Written for projects involving significant subsurface works, e.g. tunnels and caverns, the ITA Framework Checklist explains the major contractual issues that must be considered to ensure a successful project. Its goal is not to specify how each issue should be resolved, but rather to highlight the issues and give recommendations on how they should be addressed. The Checklist is compatible with, and provides guidance for reviewing, standard contracts such as FIDIC and NEC. One of the authors is a co-author of the Checklist and has contributed to previous papers describing the Checklist and its applicability to hydropower projects (Hodgkinson & Wilson 2014, Hodgkinson & Wilson 2015). The Checklist identifies a series of issues that should be contractually defined, which can be summarized as follows: •,Methods for dealing with geotechnical uncertainties should be defined in the contract •,There should be an appropriate level of geotechnical/geological investigation •,Sufficient mobilization periods should be allowed •,Sufficient time should be allowed for testing and commissioning •,Procedures for timely dispute resolution should clearly defined •,Quality and performance requirements should be clearly defined and measurable •,Provision should exist for cost consequences of unforeseen circumstances •,Dispute resolution processes should be clearly defined •,Payment processes should be fair and clearly defined •,Risk should be clearly allocated and parties should be able to bear the risks they carry •,There should be a mechanism for adjusting payments according to conditions •,The risk management approach should be clearly defined •,Admeasurement or unit price approaches are recommended •,The responsibilities of each party must be clearly defined •,The consequences of non-performance should be clearly defined •,The construction programme (schedule) should be realistic."

Jan 1, 2016

By Z. Zhongwei

The leaching kinetic of mechanically activated sphalerite with acid FeCl3 solution was investigated. The result has shown that mechanical pre-treatment has greatly enhanced the leaching process. Comparing with the original concentrate, the activated one can be leached more completely at relatively lower temperature with diluter FeCl3 concentration within a shorter leaching time. The activation energy for the activated concentrate is 27 KJ/mol, while for the original one is 52 KJ/mol.

Jan 1, 1995

By S. Rezaei

Moving ground in the mining industry is almost inevitable; however comprehensive awareness of these movements is necessary to mitigate risk. Monitoring of displacements caused by mining is a key tool for geomechanical risk management. Good monitoring may prevent unexpected failures and protect personnel and equipment. Many displacement monitoring techniques are available, however not all are reliable, accurate, and cost effective. Close range photogrammetry (CRP) is a monitoring method with high accuracy, reliability, and cost effectiveness; and has the ability to measure places that are inaccessible. In CRP, some series of photos are taken in different positions with different angles at different time intervals depending on the displacement rates. These photos are processed using PhotoModeler software and 3D models are made. By comparing 3D models, displacements can be measured. This method has been demonstrated at the Freeport-McMoRan Safford Mine in Arizona. A fault related instability in part of the pit is reactive to mining, and provided an opportunity to compare results between traditional monitoring methods and CRP. In this project, multiple series of photos were taken at different times using a Canon 5D Mark III digital camera, then PhotoModeler software used to make 3D models for each specific measurement event. Movement rates were determined by comparing different 3D models. These results are then compared to results of other monitoring methods to validate the CRP displacement method.

Feb 27, 2013

By M. C. Fuerstenau

Collophane, cryptocrystalline fluorapatite, is examined as a substrate for plant water and wastewater remediation. The attenuation of Fe3+, Cr3+, Fe2+, Cd2+, Hg2+ and Pb2+ is presented as a function of metal ion concentration, amount of collophane, initial pH and conditioning time. TCLP results for Pb2+ contaminated collophane are presented. Mechanisms of adsorption for divalent and trivalent species are suggested. Key words: Collophane, Wastewater remediation, Heavy metal removal

Jan 1, 2003

By John H. Archibald, Arthur L. Miller, George A. Savanick

As part of its continuing research on selective mining, the Bureau of Mines has made progress toward developing a tool capable of remotely extracting hard minerals through a borehole. As the successor to the successful borehole miner, which extracted soft materials through a straight borehole, this new device has several innovative features: 1) it can propel itself in and out of an 8-inch borehole; 2) it can cut hard rock with an abrasive jet and transport the cuttings to the surface; and 3) it can negotiate deviated boreholes--the feature that enables it to follow a vein and extract only the values. This paper describes the development and feature testing of the reamer/miner.

Jan 1, 1991

By G. A. Hayes

The range of gold ores tested at AMMTEC LTD over several years has resulted in the requirement to analyse for the mode of occurrence of gold in order to assist selection of ore processing options. The aim of this paper is to describe the laboratory testwork and analytical procedures which have been of benefit in detail­ing the mode of occurrence and mineral associations of gold and hence being of value in developing an optimised ore processing route. The range of ore and mineral assemblages described include relatively simple pyrite and free milling ores, more complex pyrite plus arsenopyrite gold ores and the quite com­plex associations of gold in copper bearing ores or with car­bonaceous mixed sulphide ore assemblages. The aim of this paper is to describe the laboratory testwork and analytical procedures which have been of benefit in detail­ing the mode of occurrence and mineral associations of gold and hence being of value in developing an optimised ore processing route. The range of ore and mineral assemblages described include relatively simple pyrite and free milling ores, more complex pyrite plus arsenopyrite gold ores and the quite com­plex associations of gold in copper bearing ores or with car­bonaceous mixed sulphide ore assemblages.

Jan 1, 1996