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By P. Shang

Fly ash and zeolite have strong absorption to cations since their porous microstructure. The porosities of fly ash and zeolite can also be increased by modification, which improves the absorption of the materials. This research showed that the effects of modified fly ash and zeolite to remove ammonia nitrogen from rainwater. The optimal reaction time, additive amount of materials, and pH value of the stormwater to remove ammonia nitrogen were experimentally determined. For modified fly ash, the optimal removal rate of ammonia nitrogen from 100 ml stormwater was 74.41% with the addition of 10.0g fly ash, 1.5 hours of the reaction time, and 8 of pH value. To the modified zeolite, the optimal removal rate of ammonia nitrogen was 51.22% at the reaction conditions of 2.0g of zeolite, 1 hour of reaction time, and 7 of pH value.

Jan 1, 2010

By George I. Staber

This paper is a brief review of the recent history, and the present status of, of the program designed to develop practical new processes for producing pipeline quality gas from coal. Clean synthetic pipeline-quality gas from coal became a major objective during 1971. Agreement was reached with the American Gas Association, Inc., to cosponsor a $120 million, 4-year program to be funded $80 million by Government and $40 million by American Gas Association. The program was first announced by President Nixon in his Energy Message on June 4, 1971, and a portion of the additional funds needed for the program was approved by the Congress in the December 1971 supplemental budget action. A natural gas shortage has developed in the United States and every indication is that it will worsen. Many transmission and distribution companies have already been forced to curtail sales because of the shortage. Large sums are being committed to systems needed to bring liquefied natural gas from abroad and to plants to manufacture synthetic natural gas from petroleum--also largely imported. On the other hand, there is sufficient domestic coal to supply all of the needed gas through the next century. This accelerated program is designed to provide processes for making gas from coal at costs, including security costs, below the imported alternates. As outlined in the President's Energy Message, OCR and A.G.A. have accelerated the coal gasification program to make it possible to construct and operate a "demonstration plant" (semicommercial size) after 1975. Several processes, in various stages of development as will be described later, are being supported. Concurrently, the overall program will be constantly monitored and evaluated with the assistance of an independent, unbiased contractor. C F Braun of Alhambra, California, was selected for this task and has been actively engaged in it for the last several months. Through continuing studies of the various processes and subprocesses as the pilot plant program continues, optimums will be discerned. These will serve as

Jan 1, 1972

By V. A. L. Breguncci, A. H. Martins, C. P. Ribeiro Jr.

This work presents experimental results for the recovery of gold present in alkaline cyanide solutions by using the electroelution technique of strong-base anionic polymeric resins IRA 402-CL from Rhom and Haas-Brazil Ltd. loaded with gold. This technique of electroelution of anionic polymeric resins is based on the removal of metal ions adsorbed to the resin, resulting from a potential difference applied between the cathode and anode in an electrolytic cell. The electrolytic cell was composed of an aqueous solution with suitable chemical characteristics and an anionic polymeric resin bed loaded with gold. The resin used in the experiments was loaded using synthetic alkaline gold cyanide solution with 1.0 g/L of gold. The variables studied were electroelution time, cathode current density, distance between cathode and anode and the electrolyte chemical composition. Statistical methods made it possible to establish the optimum conditions that would achieve satisfactory gold recoveries by the electroelution technique. The statistically predicted response for the resin loaded with gold by percolation was 16.5% gold recovery, 2.0 A/m2 cathode current density, 20-mm distance between electrodes and 6.0 hrs electroelution time. For the centrifuging system, the statistically predicted response was 7.41% gold recovery under the following conditions: 0.25 A/m2 cathode current density, 1-mm distance between electrodes, 3.5 hrs electroelution time and 10.9% NaOH in the electrolyte chemical composition.

Jan 1, 2002

By Kirsten Corin, Belinda McFadzean, Cyril O’Connor, Jenny Wiese

As orebodies being mined globally are of ever-decreasing grades, the focus on the concentration process has moved increasingly to the challenges faced in management of the gangue minerals. This is especially so in the flotation of platinum group minerals (PGMs), where head grades are typically in the region of 4–7 g/t. Hence, in PGM operations, a typical recovery of about 80 percent requires depressing about 85 percent of the gangue minerals. In a flotation cell the main control parameters are reagent type and dosage, froth depth and air addition rate. This paper reviews research projects carried out over an extended period aimed at optimizing the depression of gangue minerals: both those that are naturally floatable and those, such as chromite, that are entrained.  This has involved comparing the two major depressants used in the industry — guar and carboxymethylcellulose (CMC) — in the depression of the relevant gangue minerals, as well as investigating their consequent effect on the behavior of the froth phase. Because most PGM flotation plants treat the ore as if all the PGMs are associated with sulfides, copper sulfate is widely used as an activator, but this can also lead to inadvertent activation of gangue minerals. Finally, chromite levels must be kept below 3–5 percent to avoid any negative effects on smelter performance, and thus it is important to reduce chromite recovery which mainly occurs due to entrainment through, for example, froth depth control.   

By H. Douglas Dahl

In order to provide a reference framework for discussing the applicability of drilling technology to shaft construction in coal mining ventures, it is meaningful to review the history and nature of United States coal mines relative to shafting. Coal mining in the United States is characterized by shallow depths relative to most of the coal mining world. While the trend is toward deeper and deeper mines, many new mines have shaft depth less than 800 feet, and a large number of operating mines are drift or outcrop mines utilizing shafts for ventilation and portals. In the past, shafts of 200-300 feet have been commonly constructed. The use of conventionally sunk shafts at present and in the past is due in part to the shallow shaft length required, favoring conventionally sunk shafts. Another factor is that large shafts are required for multipurpose and production shafts. A third has been the practice of utilizing shaft construction contractors who had been until recently very competitive with raise drilled shafts at shallow to moderate depths. Recently, however, the costs of conventionally sunk shafts have skyrocketed. Too, more shafts are being required. Several factors are responsible among which are: (1) The increasing depth of shaft required for new mines and for shafts in advance of old mines. This significantly affects unit costs, e.g. $/yd3,of conventionally sunk shafts. Moreover, more gas is encountered in the shaft sinking operation.

Jan 1, 1976

By Ireneusz Tomza, Jolanta Lebecka

INTRODUCTION Radioactive deposits were observed in 1972 in some of the Upper Silesian coal mines. They were located mainly in the drains in galeries and on the inside surfaces of water pipes. They also caused some problems by accumulating in water pumps. It has been postulated that the deposits are produced by natural radioactive waters seeping from the rocks. Investigations were initiated to answer the following questions: - What is the composition and the amount of radioactivity in the deposits? - What radioisotopes are present in the water? - How are the radioactive deposits formed? - Do the radioactive waters also occur in other mines? - How does the radioactivity of the water depend on chemical composition? - What is the origin of the radioactive water? - Does the water and the deposits cause radiation hazards for miners? -How can the radiation hazard be reduced? METHODS OF MEASUREMENT Determination of Radium Isotopes in Water The commonly used methods of radium determination in water are either based on measurements of the radioactivity of 222Rn which is in equilibrium with 226Ra, or on the detection of alpha particles of the radium radioisotopes after chemical separation of radium from the water sample. The method based on radon activity measurements is very sensitive and does not require any chemical Separation, but it can be used for determination of 226Ra from the uranium series only, because the thorium daughter 220Rn has too short a half-life (55s to yield the required accuracy. The method developed by Goldin, 1961 [2] involved alpha-particle measurements in thin layers of RaS04 and BaS04 separated from the water. This method is not convenient for saline water and water with high barium concentration because the amount of barium carrier in this case is too large to obtain a thin layer of precipitate with sufficient activity. The Upper Silesian carboniferous waters are often saline with high barium content, so the method described by Goldin was not convenient for this case and it was necessary to change the detection system and modify the chemical preparation. The procedure developed by the authors for the determination of radium isotopes in water was as follows: - Depending on the Ba2+ content and the required sensitivity of measurement, a water sample of 200 cm3 to 3 dm3 was taken. - 10 cm3 of 0.25 M citric acid and 5 cm3 15M ammonia was added to form complex Ba2+ ions and avoid the immediate precipitation of BaSO4. (This was repeated as long as the addition of BaC12 did not form a precipitate.) - 1 cm3 of 1N solution of Pb(N03)2 as a carrier for radioactive isotopes of lead and 10 cm3 of 0.1 N BaC12 as a carrier for radium were added. - The sample was heated to the boiling point and the precipitation of RaS04, BaSO4 and PbS04 with 50% H2SO4 was carried out. - After several hours the sample was centrifuged and the precipitate was purified by washing with nitric acid and distilled water. - The precipitate was then redisolved in 20 cm3 0.125 M Na2EDTA and 3 cm3 6M ammonia and reprecipitated from the solution by dropwise addition of acetic acid to d pH of 4.5. At this value of pH, precipitation occurs only for the barium and radium sulfates, while lead and all other radioactive elements remain in the solution. The date and time of deposit precipitation was recorded. - The final barium-radium sulfate mixture was washed with distilled water and transferred to standard measurement vials. - Each vial containing a deposit had 6 cm3 of distilled water added and was then shaken vigorously. 12 cm3 of liquid gelling scintillator (INSTA-GEL UNISOLV-1 type) was then added and the vi 1 was shaken again. After a while the scintillator turns into a milky gel in which the deposit is uniformly distributed. - The standard sample of 226Ra was prepared in the same way. - The activity of the samples was measured using a liquid scintillation spectrometer. (In this case the TRICARB 3320 produced by Packard Instruments, was used). Tests run on standard radium solutions provided by Amersham Radiochemical Centre indicated that this method of measurement enables one to achieve an efficiency of almost 100% (within measurement error). For alpha particles no quenching effect was observed for the BaS04 concentration in the range up to 80 mg of BaS04 per 1 cm3 of liquid scintillator coctail (Fig. 1). This provides a sensitive determination of radium in water with high barium content and also in saline water. In saline water the solubility of barium sulfate is much higher than in

Jan 1, 1981

By Carter Brown, Kate Brown Requist, Eric Lutz, Moe Momayez, Athira Satheesh

New mining health and safety training modalities are becoming increasingly prevalent in the underground mining industry. Various digital training tools have been developed over the past several years, but the extent of their use remains unclear. Advances in gamified training by way of computerbased video games and virtual reality (VR) have opened a new avenue for mine rescue and self-escape training, but these methods have not been widely adopted across the mining industry. Mine rescue and self-escape situations require a mix of in-classroom training as well as live drills, but it is important to assess the extent of and effectiveness of new training modalities. The Mine Rescue and Self-Escape Survey (MRSES) was distributed to underground mining operations to collect perceptions of current training effectiveness as well as willingness to adopt new training methods as a complement to live rescue drills. Responses from the MRSES indicate that adoption of digital training technologies has been slow-going despite an interest in their use. Further, there appears to be no meaningful connection between more frequent escape and rescue training and feelings of emergency preparedness. With relatively little use of already developed digital training tools but an interest in their adoption, it may be time to reassess current emergency preparedness training philosophies.

May 1, 2025

By W. C. Yates

Lack of visibility from large, open-pit haulage truck cabs has become a serious safety problem. Anaconda conducted a research program to develop a low cost system to alert haulage truck operators to small vehicles present in blind areas. Several systems were developed and tested. As open-pit haulage truck size has increased, so has the size of the blind areas. Mine utility vehicles, pickup trucks, and cars are concealed within these blind areas and become safety hazards. Figure 1 shows the typical blind areas of a 136 t (150-ton) haulage truck. The areas of greatest concern are at the rear center and front right corner of the truck. In these "no visibility" areas many small vehicles have been crushed by large haulage trucks at various mine sites. In most instances, the occupants escaped the vehicle before it was destroyed. But injuries and death have occurred. In most cases, the accidents were attributed to negligence or carelessness by the small vehicle driver. The limited visibility problem still exists, though, and a solution is needed. Initial Solution The usual plane mirror on the driver's side and the convex mirror on the right front of the haulage truck allow only limited rear visibility. The typical, right side convex mirror is generally small and difficult to locate at a glance. This convex mirror also presents an image that distorts distance perception. These mirrors are of¬ten broken and immediate replacement not demanded or made. The US Bureau of Mines, in a contract with M.B. Associates, San Ramon, CA, developed an improved visibility system. It consisted of a larger, left view mirror and unique, blind area viewer. The large 230 x 690 mm (9 x 27 in.) left mirror has a small convex spot mirror near the bottom. The entire unit can be quickly replaced if broken. The large size provides a full view of driver orientation features such as the rear tire and the top of the load bed. The right 300 x 410 mm (12 x 16 in.) mirror has a convex surface and is quite rugged. The 510 mm (20 in.) spherical radius presents the operator with some distance distortion, but objects in its view are easily seen. The "blind area viewer" consists of a modified fresnel lens that allows the driver to "look around the corner." Installation of three of these lenses on the front of each haulage truck is rec¬ommended for adequate blind area coverage. This system significantly increases an operator's field of view and allows viewing of objects within about 3 m (9.8 ft of the truck. Many open-pit mines are nor mounting on smaller vehicles a long whip with a flag and lamp or both. This increases vehicle visibility to haulage truck operators. The whip is permanently attached to the mine vehicle. It can also be magnetically attached to transient vehicles. The whip system. though, is inadequate for visibility in the critical rear and right front blind areas. Alternate Solutions All of these visibility aids as¬sume an attentive driver. In addi¬tion, areas such as the rear blinc, area are not covered. Based on the inadequacy of existing visibility systems, Anaconda made a com mitment to further the truck de¬tection technology. The best ap¬proach was considered to be ar active system. It would notify the haulage truck operator that an en¬croaching vehicle is present. Sev¬eral systems were considered, in¬cluding ultrasonics and elec¬tromagnetics. For the ultrasonic system, com mercial and in-house design were looked at. Designing an( building transducers with the spe cific beam shape to match the haulage truck blind spots re¬quired a major technological breakthrough. False alarms ap¬peared difficult to prevent en¬tirely. They were caused by pass¬ing, trailing, or leading vehicles a., well as nonthreatening objects. As a result of these problems, the ul¬trasonic technique was not pur sued. An alternative, an electro

Jan 7, 1982

By Somia Gurunathan, Harry Castrantas

Hydrogen peroxide and calcium peroxide are used to improve the extraction of gold/silver and reduce cyanide consumption in heap leaching and mill leaching operations. FMC's standard grade H2O2 and FMC's Permeox™ solid peroxygen, a formulated product containing cao2, was used for the experimental work. It was found that the rates of release and the longevity of these peroxygens in ore-cyanide mixtures are significantly different. When choosing a peroxygen, one of the key factors is extrac­tion time, with the longer extraction times of heap leaching primarily favoring calcium peroxide and the shorter extraction times of mill leaching primarily favoring hydrogen peroxide. A series of laboratory leach tests showed the choice between H2O2 and CaO2 to be ore and process specific. H2O2 and/or Permeox™ improved the gold extraction in four out of five ores tested. (2 sulfide and 2 oxide). Mixtures of H2o2 and Permeox™ were also found to significantly increase the gold extraction rate. A fifth ore (sulfide) did not respond to peroxygen treatment. In addition, savings in NaCN consumption was obtained with Permeox™. Both peroxygens offer economic incentives to gold processing.

Jan 1, 1992

By Charles H. Rice

Coal liquefaction processes can be divided into two broad classifications - direct and indirect. Direct liquefaction processes, in general, slurry coal with a solvent and react it with hydrogen under elevated temperature and pressure. Indirect liquefaction is the term applied to processes in which coal is first gasified to generate a synthesis gas of carbon monoxide and hydrogen. The synthesis gas is then converted to various types of liquid fuels by the proper selection of catalyst and conditions. The first experiments on direct coal liquefaction were undertaken in 1933 by Bergius. Development and commercialization of the Bergius concept was accomplished mainly through the effort of I.G. Farben at the Ludwigshafen plant of its subsidiary, Badische Anilin und Soda Fabrik A.G, under the leadership of M. Pier (1). A demonstration plat having an annual capacity of 100,000 tons of coal was built in Leuna, Saxony in 1927. Six years later, Imperial Chemicals Industries, Ltd., commissioned a plant in Billingham, England (2).

Jan 1, 1981

By K. A. Natarajan, K. Karthiga Devi

"A biodegradable flocculant was produced during growth of Bacillus megaterium. The major component of the bioflocculant was found to be a polysaccharide composed of some proteins. Fourier transform infrared (FTIR) spectra analysis revealed the presence of carboxyl and hydroxyl groups in the bioflocculant, and thermal characterization by differential scanning calorimetry (DSC) showed the transition and crystalline melting point at 90-105°C. The effects of bioflocculant dosage and pH on the flocculation of mineral suspensions were evaluated. The bioflocculant exhibited good flocculating capability on mineral suspensions and achieved flocculating efficiencies of 90 percent for kaolinite and 85 percent for hematite suspensions at a dosage of only 5 mL/L. The maximum arsenite removal was found to be 90 percent at a bioflocculant dosage of 2 g/L, which is better than traditional chemical flocculants. This study demonstrates that microbial bioflocculants have potential for application in environmental cleanup, such as in the flocculation of mineral fines and in the remediation of solutions that contain toxic heavy metals.IntroductionBioflocculants are biopolymers produced by microorganisms during their growths. The predominant components of a bioflocculant are extracellular polymeric substances such as glycoprotein, protein, cellulose and nucleic acid (Salehizadeh and Shojaosadati, 2001). Bioflocculants have gained much wider attention due to their biodegradability and safety. In recent years, the study of bioflocculants has attracted wide attention. Bioflocculants can be nontoxic, harmless and eco-friendly. Normally, they exhibit different functional properties, serving as stabilizer, dispersant or thickener. Among these properties, flocculation is of major interest. For example, Phormidium sp. strain J-1, isolated from a drainage channel, exhibited excellent flocculation activity for bentonite particles in liquid suspension (Fattom and Shilo, 1984). A protein bioflocculant produced by Rhodococcus erythropolis could efficiently flocculate suspended solids while losing their flocculating activity by enzymatic digestion (Kurane et al., 1994, Takeda et al., 1991). Klebsiella sp. strain S11 screened from activated sludge was unable to flocculate without the addition of calcium chloride (Dermlim et al., 1999)."

Jan 1, 2015

By Robert E. Ennis

Several important economic variables must be studied when planning a major underground conveyor haulage system if the objective of tow cost haulage is to be met. The purpose of this paper is to call these economic factors to the attention of the mine planner and encourage their serious consideration. To illustrate, the importance that some of these variables hole for the mine operator who wishes to reduce costs and maintain sound belt haulage practice, these areas of discussion will be covered. 1. Selection of economical flite lengths. 2. Economics of wire rope frame vs. rigid frame conveyors. 3. How idler design affects conveyor costs. 4. Economies in drive design. 5. How speed affects conveyor cost. 6. P.V.C. vs. Neoprene belt - what savings? 7. Maintaining, conveyor investment through maintenance. 8. Miscellaneous factors affecting the cost of belt haulage.

Jan 1, 1967

By Dale S. Preece

INTRODUCTION The Waste Isolation Pilot Plant (WIPP) is a research facility located in Southeastern New Mexico. The WIPP is being developed by the U. S. Department of Energy (DOE) to demonstrate the safe storage of defense-related radioactive wastes in bedded salt. Since the WIPP is a research facility. a number of large-scale in situ experiments have been planned and are currently under construction (Munson, 1983). These experiments were designed to study various aspects of nuclear waste storage in bedded salt such as mechanical responses and creep closures of drifts, thermal response due to heated canisters, and thermally-induced fluid migration. One purpose of the WIPP is to develop and demonstrate a general predictive structural analysis capability for a bedded salt repository. One set of experiments, called the Geomechanical Evaluation (Room G), is heavily instrumented to study the creep around several rooms with different configuration. A Layout of the Geomechanical Evaluation room is shown in Figure 1. One portion of it consists of wedge shaped pillar where two drifts intersect at an angle of 7.5 degrees. The pillar can be seen at the bottom of Figure 1. One purpose of the wedge pillar experiment is to study progressive pillar failure in the tip region. Another is to determine how nonuniform pillar thickness affects creep closure of the drifts. An important aspect of the experiments is the correlation between experimental data and corresponding pretest finite element analyses of the site. The finite element simulations serve two purposes. First, the computer simulations aid in understanding the experimental results by providing calculated stress. strain and displacement fields that cannot be measured directly. Second, comparison of calculated and measured drift closures serves as a method for validating or improving the finite element models and the constitutive models employed. The wedge pillar geometry required a 3-D finite element creep calculation. Results from these 3-D calculations will be presented in this paper. A 2-D plane strain double drift model which approximates the wedge pillar geometry at a slice perpendicular to the drift has also been performed and will be compared to the 3-D results in this paper. FINITE ELEMENT COMPUTER PROGRAMS Two finite element computer programs. JAC (Biffle, 1984) and JAC3D (Biffle. 1986), were used. JAC is a 2-D finite element program developed for quasi-static analysis of non-linear solids. It employs the conjugate gradient iterative technique to obtain a solution Spatial integration is performed using a single gauss point in each four node quadrilateral element. An hourglass viscosity technique is used to control the zero energy modes that occur with single point integration. The single point integration combined with the explicit nature of the program and exploitation of CRAY-1 computer architecture results in very efficient execution J4: results were compared to results obtained with eight other 2-D structural creep computer codes in the second WIPP benchmark exercise (Morgan, 1981) JAC3D was derived from JAC to treat 3-D finite element models and has many of the same characteristics including single point integration and hourglass viscosity. These characteristics have made 3-D creep analyses more reasonable by significantly reducing computation time The first exercise of the creep capability

Jan 1, 1986

By Dapeng Zhang, Brij Moudgil, Shibo Zheng, Chunhui Feng, Wenyi Song, Wending Xiao, Patrick Zhang

The separation of dolomite from phosphate is the most challenging problem in phosphate mineral processing. More than 50 percent of the phosphate reserve in Florida contains too much dolomite to process using current industry practice. The current way of producing phosphoric acid requires less than one percent MgO in the phosphate concentrate as the feed material. The Florida Industrial and Phosphate Research (FIPR) Institute has collaborated with worldwide experts in the field to address this issue. As a result, the industry is now offered three feasible options. Option I entails three methods for reducing MgO content in the concentrate from the Crago process, including adding a dolomite depressant in the rougher flotation step, dolomite flotation of the cleaner concentrate, and scrubbing the cleaner concentrate in quartz sand, which could reduce MgO content in the final concentrate by 20 to 40 percent. Option II involves the crushing and grinding of high-dolomite phosphate pebbles followed by dolomite flotation at slightly acidic pH using a new collector that does not require phosphoric acid as a phosphate depressant, achieving a final concentrate analyzing less than 0.9 percent MgO at about 87 percent P2O5 recovery. Option III is a gravity separation technique using an innovative separation jig, and its full potential remains to be demonstrated.

By S. S. Hsieh

As part of a program to address the long-term need to ensure adequate supplies of domestic phosphate raw materials, the Tennessee Valley Authority (TVA) has conducted considerable research on dolomitic phosphate rock beneficiation. The discovery that alkyl diphosphonic acids act as phosphate mineral depressants led to the development of several processes involving a carbonate flotation step for the effective beneficiation of these problem ores. Most recently, a partial deoiling process in which the reagent consumption is greatly reduced has been developed. The reduction is achieved because part of the initial feed is beneficiated without passing through the separate carbonate flotation step. Using this method, a phosphate product containing 30.6-30.9% P2O5 and 0.7-0.8% MgO was obtained from a Florida dolomitic phosphate pebble that contained about 23% P2O5 and 2.0% MgO. The P205 recovery was 80.8-86.5%.

Jan 1, 1991

By Talaat A. B. Lawendy, Guerry H. McClellan

Batch laboratory-scale flotation tests were conducted on some carbonate-containing sedimentary phosphate ores to obtain concentrates suitable for chemical processing. The phosphate samples tested contained both calcite and dolomite. The calcitic phosphate ore was from Israel and the dolomitic ores were from China, India, Egypt, and Pakistan. The phosphate sample from Pakistan also contained a significant amount of silica. Several flotation techniques that have been proposed for upgrading carbonate -containing phosphate ores and concentrates were tested. Two processes gave better results compared with the other processes tested: a modified IMC anionic process and a. phosphoric acid depression process. At the best flotation conditions tested, phosphate concentrates assaying 32.5%-36.5% P205, 44.0%-51.0% CaO, and 0.6%-1.0% MgO were obtained at 69%-76% overall P205 recoveries from run-of-mine (ROM) ores containing 16.0%-25.0% P205, 40.0%-52.0% CaO, and 1.0%-11.6% MgO.

Jan 1, 1993

By A. D. Hammond

Until recently, the volumes delivered and unloaded from heap-leaching pads have been estimated by manual methods at Wharf Resources' South Dakota operation. In August 1989, the Engineering Department acquired the MINE-SURVEY software developed by GEM COM Services, Inc. of Vancouver, Canada, as part of a mine planning/grade control software package. The implementation of MINE-SURVEY for pad control resulted in a cost-effective tool to provide accurate and prompt estimates of ore delivered or unloaded from the pads. Volumes can be calculated in a matter of minutes as opposed to hours or days. Another benefit is the ease of comparison of as-build and designed volumes, plotting cross-section in any orientation, importing into AUTOCAD, validation of ore densities and truck factors. This paper describes the application of the MINE-SURVEY systems in heap leaching pad volumetrics.

Jan 1, 1991

By V. O. Tenorio, K. P. Hague

Mines are constantly trying to improve production processes by optimizing work cycles, reducing operational delays, minimizing costs, controlling equipment utilization, and preventing occupational hazards. Currently, advances in digital mining technologies within the concepts of 5G and Industrial Touch Internet are allowing better control and monitoring of underground mining equipment, with centralized control systems that unify data analysis for proper decision making. A wireless 5G-based platform for controlling and monitoring systems for underground mine operations can be designed over a comprehensive layout in where bandwidth selection, sensor location, Wi-Fi repeaters and power distribution are put together to build a successful supervisory system. In order to confirm the validation of these concepts, a case study is proposed for a medium-sized underground mine with a multi-level network which includes tablets managed by production supervisors, along with other data collection devices installed in semi-autonomous mining equipment, collecting, and distributing data in near-real time, with interconnection to support and services areas on the surface.

Mar 2, 2022

By A. E. Petersen

The United States Bureau of Mines conducted research on high-intensity magnetic separation methods for upgrading tantalum concentrates from Alabama. Process flow sheets were studied to maximize tin removal and minimize tantalum loss. Using laboratory-scale equipment, concentrates were produced which contained over 65% combined Ta2O5-Nb2O5. Up to 78% of the tin was recovered as a byproduct, containing about 70% SnO2. Only 1 to 5% of the tantalum reported with the tin byproduct. The producer subsequently incorporated high-intensity magnetic separation into its process flow sheet to provide concentrates for the National Defense Stockpile.

Jan 1, 1993

By G. Van Weert

Oxygen transfer from dispersed air in agitated slurries represents an important unit operation in the minerals processing industry. Gold leaching, iron precipitation and especially bio-oxidation come readily to mind. However, little engineering information is available on the variations between impeller types on such O2 transfer. Engineering literature on the subject, i.e. information which relates slurry density and energy input to O2 transfer, is reviewed and compared to current experimental work. The latter was carried out with minus 50 micron sand/water slurries in a 64 I laboratory tank of standard proportions, equipped with a variable speed drive with torque meter. A standard Rushton turbine and a specifically sized A315 were compared. Results confirm an increasing impairment of O2 transfer with increasing slurry density for both impellers. The volumetric mass transfer coefficient, k,a, in a 10 vol. % slurry shows a decrease of 30% compared to the k,a in water. At 20 vol. % sand this decrease is near 50 %, depending somewhat on superficial gas velocities. Power efficiency, i.e. the kg O2/kWh, decreases about 25 % more rapidly for the Rushton turbine than for the A315 with the addition of sand. However, for equivalent O2 transfer the Rushton turbine was more efficient in this work. Experimental results are presented and their significance for slurry bio-oxidation discussed.

Jan 1, 1996