Search Documents

By Matthew Holden, Jason Emery, Matthew Benjamin, Peter Craig

Over the last 25 years, through adversity and a growing level of risk mitigation, Australian Longwall mines have been using higher and higher capacity cable bolts for secondary support. A recent survey revealed 85% of Australian longwall cable bolts are a 70 tonne capacity pre-tensioned hollow post groutable system. Mines avoiding pumping cementitious grouts are sometimes using pre-tensioned resin encapsulated solid cable bolts of over 90 tonne capacity. Results from various industry and University research has led to manufacturer's fine tuning the construction of their cable bolts. A review of this research explains how smooth wire construction is now dominant. Aggressive shear has also led to some mines using 8 tonnes pre-tension instead of the previous common 20 tonnes pre-tension. Lower pre-tension acceptance is now leading to a greater use of threaded cable bolt end fittings, these are tensioned to 8 tonne using the drill motor torque. A recent development is the TEX cable bolt, a slim line threaded end fitting that fits with the drill hole like a standard rock bolt. This cable is being installed at the coal face as primary support off bolter-miners. The laboratory and in-situ testing of the TEX cable is presented. In a bid to improve gateroad development rates and ground support in the last several years, a 100 tonne tensile capacity Goliath cable system was developed. The latest development in resin anchored solid cables is a 110 tonne capacity cable bolt, with laboratory and in-situ testing results presented.

Jun 25, 2024

By Harry E. Holzschuh

The two Agrico Chemical Company self unloading Phosphate Rock barges put into service in 1976 operate at over their design carrying capacity. Each barge now carries 26,000 short ton instead of the planned 25,000 short ton at 32 foot of draft. They are in service between Agrico's Tampa, Florida Terminal and the Donaldsonville, Louisiana Fertilizer Complex, moving a total of 2,500,000 short tons each year. This was the first use of the Parascrew design for a marine application. The barges, with their 3,000 STPH unloading and matching loading rate is the vital link in this total transportation and handling system. Rock is moved from the Pierce, Florida Mines through the Tampa Big Bend Terminal, across the Gulf of Mexico, and up the Mississippi to the Agrico Faustina Phosphoric Acid Plant 30 miles below Bacon Rouge, Louisiana. The trip from the terminal to the plant is 600 miles and the turnaround time is just about six days. Seven years operating experience has proved that the system works. Fittingly, the two barges are named the "Faustina" and the "Pierce". They have a beam of 85-feet and are 495 feet long. folded depth is 43 feet, with a ballasted light draft of 19 feet and a typical loaded draft of 32 feet. There are four separated holds with the machinery room located forward of 101 hold. Each hold is 51 feet wide at the coamings with an 18 foot straight section on the side, and a 24 foot long, 600 slope sheet for a 23 foot clear reclaim width at the bottom. The Parascrew in each of the holds is 37 feet long between the bearings and is supported on a 14 inch diameter hollow shaft. The screw pitch varies and the diameter of the flights step up from l'-9" diameter to 2'-6" diameter on the discharge end. By sloping the centerline, a minimum four inch bottom clearance is maintained across the hold. Some bending of the tube does take place.

Jan 1, 1983

By R. S. Olsen

The Bureau of Mines is engaged in research to recover aluminum from kaolinitic clay and other nonbauxitic domestic resources. As part of this work, calcined Georgia kaolin was prepared by a process that prevents slime formation and makes the leached residues fast settling and easy to filter. The leaching reaction for calcined kaolinitic clay in a 5 pct stoichiometric excess of boiling 23-to 26-pct HCI was found to be zero order. Rate constants based on the weight-fraction of aluminum in the clay were found to be between 0.095 and 0.15 min-1 depending on feed preparation and leaching parameters. The mathematical relationships between aluminum extraction and time for both batch and continuous, stirred-tank operations were derived and verified experimentally.

Jan 1, 1983

By Baruch Grinbaum, Edwin Bwalya, Robert Minango, Deogratias Bukunkwe

Solvent extraction of Cu from acidic leach, using an oxime extractant, e.g. LIX 984 or Acorga is a well known and established process. As the process advances, Cu is extracted to the solvent and is replaced by H+ in the aqueous phase. At high acidity the distribution coefficient of Cu between the solvent and the aqueous phase decreases, and eventually relatively high concentration of Cu remains in the raffinate, regardless of the number of equilibrium stages. In many processes this phenomenon is immaterial, as the raffinate returns to the leaching and no copper is lost. But when the copper raffinate is used for processing of additional metals, minimal concentration of Cu in the raffinate is essential for an effective process. The recommended process includes three steps: 1. SX-1: Standard extraction of Cu with oxime. The resulting raffinate contains about 6-8 g/L Cu and 50 g/L H2SO4. 2. SX-N: Extraction of most of the acid into an amine-based solvent, and stripping it with a base, e.g. aqueous ammonia, caustic, Na2CO3 etc. The extraction removes no metals, but decreases the concentration of acid from 50 to 10 g/L. 3. SX-2: Extraction of the neutralized solution once more with oxime, to remove the residual copper. Such a process enables one to get down to 100-150 ppm Cu in the final raffinate. This process was developed and tested in the R&D facility in Chambishi Metals plc, by a joint team of Chambishi & Bateman, using Cognis LIX 984N and real solutions from the plant. In the laboratory scale, the equilibrium isotherms for SX-1, SX-2 and the neutralization were developed. Computer simulation of the results showed that the optimal configuration is two stages for both extraction and stripping for SX-1. For SX-2 2-3 stages of extraction and two stages of stripping are needed. The stripping stages may be common to both SX-1 and SX-2. The neutralization of SX1 raffinate was tested both by conventional precipitation with lime and by extraction with Alamine based solvent extraction and stripping with a solution of 10% NaOH. For the SX-N only one stage was needed for both the extraction and stripping. The bench scale pilot plant was run for both SX-1 & SX-2. The results were very similar to the simulated ones, and the configuration of 2E2S for SX-1 and 3E=2S for SX-2 delivered the required raffinate with less than 150 ppm Cu.

Jan 1, 2005

By E. Tarshizi

Research was conducted to develop a simulation and animation model for an aggregate (sand and gravel) mine operation. This model can assist the mine engineers to investigate and consider the various options of haulage systems which are critical for the mine operation productivity. Discrete-event system simulation studies can be used for the typical haulage issues that mining engineers are often faced with: using truck fleet haulage, conveyor belts, or a combination of both. The simulation has identified some strengths and weaknesses in the current loading and hauling fleet of the mine, and has resulted in some suggestions for improving the system. The simulation includes animation of the operation. Animation is needed to enhance the benefit of any mine simulation model. By combining the visual power of animation with a mine simulation, a complete picture of the mine system is obtained and displayed. The mine is modeled and animated using GPSS/H® and PROOF Professional®, both of which have shown great applicability in discrete mining simulation projects.

Feb 23, 2014

By J. P. Powell

The importance of recovering more oil from known petroleum reserves has placed increasing emphasis on the application of secondary producing methods. One of the most successful of these has been the injection of water into depleted oil reservoirs. Water flooding is now recognized as a valuable conservation measure that not only prevents the premature abandonment of stripper leases but also is credited with reviving operations in many oil-producing areas previously abandoned. The injection of water into an oil-producing zone as a method of stimulating oil production was legalized in Kansas in 1935. Although water flooding had been attempted previously in other parts of the State, 1935 marked its introduction into the shallow stripper area of eastern Kansas. Two years later flooding was begun in the five-county area that is described in this report. Figure 1 is a map showing the water-flood areas in Allen, Bourbon, Crawford, Labette, and Neosho Counties, Kans. To July 1, 1955, the 77 projects in the oil fields of these 5 counties are credited with recovering an estimated 12.3 million barrels of oil that would not have been produced by methods used previously. These projects are expected ultimately to cover 10,700 acres, of which only 5,400 has been flooded to date.

Jan 1, 1957

By Raymond A. Povirk, David L. Bartsch, James V. Hamel

The Ohio Valley Coal Company is expand- ing its Powhatan Mine No. 6 near Alledonia, Ohio. This expansion includes a 20 rn (65 ft) height in crease for an existing 44 rn (145 ft) coal refuse - slurry impoundment dam. The dam is being raised by upstream construction which will involve placement of more than 15 m (50 ft) of new fill over an existing reinforced concrete decant structure. This structure is 0.76 rn (30 in.) square inside and its deepest portion was under 24 m (80 ft) of coal slurry and water. The top of the decant structure was not strong enough to safely support the new fill and, because of the slurry and water, there was no way to reinforce the structure from the outside. The structure was reinforced frorn the inside by sliplining 73 m (240 ft) of 0.66 rn (26 in.) OD, 0.53 rn (21 in.) ID polyethylene pipe, grouting around this pipe with cellular concrete. and constructing a steel-fiber reinforced micro-silica concrete transition at the upslope end of the pipe. Rehabilition of the decant structure was completed in August 1989. The rehabilitated decant structure is now performing satisfactorily as fill is being placed over it for raising the dam.

Jan 1, 1991

[Participating Scholarship Colleges and Universities 138 AmerCable 130 Andalex Resources, Inc 82 Behre Dolbear & Company, Inc 83 Commercial Testing & Engineering Company 86 Continental Conveyor & Equipment Co 87 Crane & Hoist Sales Inc 12 Crested Butte Mountain Resort 88 Dakota Coal Company 89 Empire Machinery Company 90 Energy West Mining Company 91 GH Hensley Industries 92, 93 Hitachi Construction Machinery (America) Corp 96 Ingersoll-Rand Company 131]

Jan 1, 1998

By Luiz Guilherme Isfer Maciel, Marilena Cardu, Fernando Golin, Jacopo Seccatore

"This study applies the concept of the difficulty of tunnelling by drilling an blasting (D&B) to selectedcases from the Brazilian tunnelling industry. The difficulty of tunnelling has been defined in previouswork as the difference of specific drilling (SD) from industrial average trend: easier when SD is lowerthan average, more difficult when SD is higher than average. Results show how tunnels in Brazil tend tobe unrealistically more difficult to excavate when compared to similar excavations in similar geologies elsewhere in the world. It is thoroughly discussed the reason for such anomaly (up to +90% SD compared to industrial average). One reason could be the lack of options amongst explosive products on the brazilian market. Only ANFO and explosive emulsions are available in Brazil. Lacking high VOD(Velocity of Detonation) products might result in reducing the maximum allowable burden amongstproduction holes, therefore increasing SD. Another reason could be found in the tunnel designers’tendency of choosing D&B as excavation method for weak rock masses that would benefit from mechanical excavation. The reason to this tendency can be partly found in the lack of availability of mechanical excavation machines in the brazilian market. Blasting in weak rock masses requires more careful smooth blasting, hence a closer spacing amongst contour holes to avoid costly overbreak: this increases SD. A third reason could be found in the lack of skill amongst the local workforce. When dealing with unskilled workers the blast designers naturally tend to avoid highly optimized blasts thatrequire very skilled workforce to be realized on field. The designer in this case tends to leave the blastdesign simpler but more robust, at the cost of higher specific drilling and specific charge. The conclusions suggest that this unrealistic difficulty when compared to world average could be a simptom of the so-called “Brazilian cost” within the tunnelling industry."

Jan 1, 2017

By Yi Luo, FENGLIANG WU

Explosions and fires originated from longwall gob due to the formation of methane-air mixture have been a severe threat to coal miner’s lives.Many numerical studies on coalmine fire and explosion hazards have focused on the airflow in roadways and mine gobs. However, most of these studies isolate the gob from its surrounding roadways, and the network analysis and the CFD method are applied independently tomodel the two classes of airflows. This approach greatly limits the ability of simulatingmine ventilation flow, especially unable to consider the effects of gob boundary conditions on air exchange between the gob and the surrounding airways. An innovative finite tube method (FTM) is developed to couple the one-dimensional mine ventilation network (MVN) and the 2D/3D gob flow field (GFF). In FTM, GFF is discretized into a finite number of flow tubes each of which is formed by any two adjacent stream lines. These tubes, representing the gob’s field flow, connect the MVN into a new coupling network. To solve the coupling model betweenMVN and GFF, an iterative solution technique is developed inwhich the MVN analysis is used to evaluate the boundary pressures for GFF simulation and in turn the FTM feeds the GFF results back to the coupling network. Based on the FTM approach, a model for gas migration in gob has been established for delineating the hazard zones of explosive methane concentration and spontaneous combustion. A computer program is developed to implement the FTM simulation. An illustrative example with five flow tubes representing the GFF is created to verify the stability and convergence of the FTM solution process. A simulation example also indicates that the accuracy of FTM is improved by 12% compared with previous method. Results of an application case show that the program is capable of quantitatively evaluating the gob’s risk zones prone for spontaneous combustion and gas explosion as well as performing risk analysis for various ventilation scenarios.

Jul 10, 2020

By Manuel Pérez-Tello, María M. Salazar-Campoy, Óscar Rodríguez-Hoyos

A common problem during the operation of the Kaldo furnace is the formation of accretions and blockage of the injection lance due to splashing. This typically occurs during the processing of anodic slimes for the production of doré. In this work, a 1:7 scale model of the Kaldo furnace currently in operation at Metalúrgica de Cobre, Mexico, was built in transparent acrylic plastic. Water and colored oil were used to represent the doré and slag, respectively. The model was used to determine the maximum penetration distance of the injection lance under a variety of experimental conditions at room temperature. Input variables included two types of oil, rotation speed, inclination angle, and the fraction of the vessel volume occupied by the liquids (φ). A mathematical correlation for the dimensionless maximum penetration distance of the injection lance d ∗ as a function of φ, the Froude number of the liquids in the vessel (Fr), the Reynolds number of the gas jet (Re), and the oil-to-water kinematic viscosity ratio (η)was developed. Reasonable agreement between the calculated and experimental values was obtained. Digital images showed that splashing was caused by the turbulent gas jet impinging upon the free surface of the liquids, whereas the cascade effect was negligible. The results of this work have been useful to improve the operation of an industrial Kaldo reactor.

By Madan M. Singh

The use of fluid jets for industrial purposes has received considerable attention in recent years. The convening of The First International Symposium on Jet Cutting Technology1 in April 1972 devoted solely to this subject probably heralds the coming of age of this new science. It is interesting to note the emphasis in the conference towards the cutting of rocks, minerals, and soils, indicating the attention being devoted to excavation and mining. The use of water streams for mining of minerals is not new. It is well over a century since its first application.2 But its use in coal mining has been somewhat restricted over the years. In the past 30 years the U.S.S.R. has successfully used water for coal extraction, and these techniques have been examined in 15 countries in recent years. In the United States research on various aspects of the method has been conducted for over 15 years,3 but success stories from the Soviet Union,4 JapanS and Canada6 in the last 5 years has spurred greater enthusiasm. IIT Research Institute has done pioneering work in the 7-12 13-15 area of water jetting for rock and coal breakage. It has just completed an experimental study involving mining of coal with a 50,000-psi and 100,000-psi jet in a stripping operation14 under the sponsorhip of the U. S. Bureau of Mines (Contract No. HOll1789). As a complementary program the Bureau initiated a project to design a water-jet coal-mining machine (Contract No. H0133119). IITRI was a successful bidder on this competitive procurement.

Jan 1, 1974

By S. Lohmeier, D. Gallhofer, B. G. Lottermoser, T. Schirmer

At a time of resource consumption, it is important to study the chemical composition of mining and metallurgical wastes to prevent the dissipative loss of metals and metalloids from the mining value chain. In particular, the recovery of critical elements from wastes is an option to increase the resources of such materials that are economically significant and have an overall supply risk. In this paper we report on the chemical composition, in particular the critical element content, of granulated slag originating from historical smelting activities in the Tsumeb area, Namibia. Laboratory-based inductively coupled plasma–mass spectrometry (ICP-MS) and X-ray fluorescence (XRF) analyses as well as portable X-ray fluorescence (pXRF) demonstrate that the slags are on average enriched in base metals (Cu 0.7 wt%, Pb 2.7 wt%, Zn 4.7 wt%), trace metals and metalloids (Cd approx. 50 mg/kg, Mo approx. 910 mg/kg) as well as critical elements (As approx. 6300 mg/kg, Bi approx. 3 mg/kg, Co approx. 200 mg/kg, Ga approx. 100 mg/kg, In approx. 9 mg/kg, Sb approx. 470 mg/kg). While metals and metalloids such as As, Mo and Pb can be determined reliably using pXRF instruments, the technique has inherent limitations in evaluating the contents of certain critical elements (Ga, Sb). However, there are positive correlations between the As, Mo, and Pb contents determined by pXRF and the Ga and Sb contents obtained through ICP-MS and XRF. Thus, quantitative pXRF analysis for As, Mo, and Pb allows calculation of Ga and Sb abundances in the slags. This work demonstrates that pXRF analysers are a valuable tool to screen smelting slags for their chemical composition and to predict the likely contents of critical elements.

Mar 1, 2021

By J. O. Marsden

Over the past decade, the copper industry has been through an extended downturn followed by several years of sustained high copper price. A number of new technologies have been developed over the past 10 years aimed at reducing costs by decreasing energy, labor, and materials requirements, or by increasing metal recovery. The energy intensity of different copper extraction routes and different technologies varies considerably and this has a significant impact on the economics of extraction. Based on average ore grade materials, the total energy for the extraction of copper typically varies from about 24-26 MJ/lb (6.7-7.2 kWh/lb) for the mining, crushing, milling, flotation, smelting and refining route to about 14¬16 MJ/lb (3.9-4.5 kWh/lb) for the mining, crushing, heap leaching, solution extraction and electrowinning route. New technology developments are providing opportunities for significant reductions in energy consumption for copper extraction, representing energy savings up to 20-25% in some cases. In addition, in many cases, these new technologies have introduced environmental, safety and health benefits over the incumbent technology. Some of these technology developments have been more successful than others, and many hold great promise for application in the coming decade. This paper provides an overview of the technology developments that have occurred, and explores the underlying reasons behind successful implementation.

Jan 1, 2007

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

Separation of dolomite from phosphate is the most challenging problem in phosphate mineral processing. Over 50% of the future phosphate reserve in Florida contains too much dolomite to process using the current industry practice. The current phosphoric acid production practice requires less than 1% MgO in the phosphate concentrate as the feed material. The Florida Industrial and Phosphate (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 1 offers 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. These methods could reduce MgO content in the final concentrate by 20–40%. Option 2 involves 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% MgO at about 87% P2O5 recovery. Option 3 is a gravity separation technique using an innovative separation jig, and its full potential remains to be demonstrated.

[Aamot, Clav Crone, (Member '29) Met., Det Norske Aktieselskab for Elektrokemisk, Industri, Inc., Oslo, Norway. Aamot, 0, Crone, (Member '29) c/o Elektrokemisk, Radhusgaten 23, Oslo, Norway. Aase, Glenn D., (Junior Member '40) Jr. Met., American Smelt. & Refin. Co., Selby, Calif. Abadie, Henry G., (Junior Member '43) Div. Petr. Engr., Union Pacific R.R. Co., 422 W. 6th St., Los Angeles 14, Calif. Abbott, Clarence E., (Member '04) Civil Engr., V.P., Tennessee Coal, Iron & R.R. Co., 1542 Brown Marx Bldg., Birmingham, Ala. Abbott, Fred E., Jr., (Member '44) Petr. Engr., c/o Arabian American Oil Co., A.P.O. 816, c/o Postmaster, New York, N. Y. Abel, Raymond L., (Member '28) Chem. Engr., 0. 0., Pittsburgh Chem. Warfare Procurement Dist., American Bank Bldg., Pittsburgh, Pa.]

Jan 1, 1946

By W. M. Myers

[Utilization of small stone is recognized as one of the important problems of the lime industry. The term ?small stone' is used to designate material that is too small for calcination in tie shaft kilo. The maximum dimension A such stone rosy be placed at four inches. The introduction of any considerable quantity of stone having demensions lee: than this into the charge of a kiln impedes the draft and readers successful operation of the kiln difficult on impossible. Attempts design a shaft kiln which would calcine stall stone satisfactorily have not been successful to date. Thousands of tons of this small stone arc produced daily, and while part of it finds a market, for it has many uses, the profit is usually small an competition is teen and prices are low. Production of the material is, however, so greatly in excess of market that surplus of hundreds or possibly thousands of Jose axe place daily in stock piles and classified as waste. in a few instances a fortunate combination of circumstances, such as a program of or other work requiring large quantities of concrete aggregate, may mate it possible some quarries to market this stone later at a profit, teat in general the chances A profitable utilization are slight, and the quarry operator has the expense of handling this waste stone and providing a storage site for it. The brittle nature of limestone makes it impossible to quarry the stone and crush, it to size suitable for the shaft kilns without producing an excess of undesirable fines. Elimination of fines by changes in quarrying or crushing practices cannot lee expected. On the contrary, the adoption of more modern methods of quarrying in recent years has increased the proportion of small material. The tendency is to blast the roan with large charges of explosives and crush it with mechanical crushers, which produces more fines than the older methods of careful blasting with small char yes and sledging the rock by hand.. We increasing use of underground mining of limestone has also increased the production of small stone, as blasting in the confined space of an underground stope shatters the rocs more completely than quarrying in the open. The use of the very friable limestones, such as the limestone of Florida for the manufacture of lime adds a new factor to the calcination of small stone. These friable lime limestones have a low tensile strength and crumble in the kiln, causing trouble. Successful calcination of these stones is accomplished in specially]

Jan 1, 1926

By Eugene T. Sheridan

The United States has extensive reserves of peat. Peat and muck lands are estimated at approximately 79 million acres, with reserves of 13.8 billion net tons. Within the past decade domestic production of peat has increased 2-1/2 times, and apparent consumption has tripled. During the same period imports have increased greatly, and in 1955 they almost equaled domestic production. The substantial increase in the demand for peat in the United States was due chiefly to expanding markets in agriculture and horticulture, for the reason that in this country peat is used primarily as a source of organic matter to improve the physical condition of mineral soils and as a mulch for shrubbery. A valuable constituent of any soil is its organic matter, and its depletion will injure cultivated soils and subject them to greater erosion.

Jan 1, 1957

By Amar Prakash, V. M. S. R. Murthy, Ajoy Kumar Singh, Avinash Paul, John Loui Porathur

Weak, layered, and fragile rock mass formation, if not supported properly, is subject to roof failure thereby affecting safety and productivity in underground coal mines. Though Central Mining Research Institute – Indian School of Mines Rock Mass Rating (CMRI-ISM RMR) based well-defined support design guidelines are established, still occurrence of roof failures in underground coal mines is a real matter of concern for mining engineers and researchers. Numerical modeling techniques are successfully used by several researchers by simulating rock mass condition for stability assessment of mine openings. The analysis becomes crucial in case of weak and fragile rock formation. The present research envelops the determination of 31 cases of rock load by CMRI-ISM RMR under different geo-mining conditions followed by the development of modified Rock Mass Rating (RMR), i.e., rock mass rating dynamic ( RMRdyn) by incorporating P-wave velocity as a new parameter. The rock load determined using RMRdyn and numerical models was correlated and found in close agreement. In addition, the deviation in rock load determined by all the three approaches, i.e., CMRI-ISM RMR, numerical modeling, and RMRdyn, was compared with the actual field data. The percentage deviation obtained in RMRdyn and numerical modeling is less compared to CMRI-ISM RMR.

Sep 1, 2023

By L. J. Froisland

The U.S. Bureau of Reclamation requested U.S. Bureau of Mines (USBM) assistance in developing design data for moving lead-zinc-gold tailings from their current location without disrupting the existing chemically stable conditions. This report presents results of USBM work in determining (1) the minimum required time to air dry the tailings to approximately 20 pet moisture under various drying conditions both in the laboratory and in the field, (2) the degree of oxidation or reduction that occurs during drying, (3) the effect of lime or cement addition before drying, and (4) the likely equilibrium conditions of the dried tailings after deposition at the new location. The limited number of tests performed by the USBM in the available time frame established trends in oxidation levels but did not provide absolute statistical validity of data values. All data from drying and oxidation testing are included in appendices to this report.

Jan 1, 2010