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By Y. C. Guan

The leaching kinetics of chalcopyrite in ammonium iodide solutions with iodine has been studied using rotating disc. The variables studied J include the concentrations of lixiviants, rotation speed, pH of the solution, reaction temperature, and reaction product layer. The leaching rate was found to be independent with the disc rotating speed. The apparent activation energy was measured to be about 50 kJ/mole from 15°C to 35°C, and 30.3 kJ/mole from 35°C to 60°C. Experimental findings were described by a kinetic model, and the overall rate was found to be: (formula)

Jan 1, 1997

By John E. Miller

In recent years there has been increasing concern in the mining industry over the availability and high cost of diesel fuel to support mining operations. Unit Rig & Equipment Co. has been deeply involved with the development and testing of the newest designs for trolley assist. Beginning in dune 1980, Unit Rig modified and commissioned six 156 metric ton (170 short ton) trolley assist haulage trucks at two mine locations in South Africa. Operating haulage trucks with General Electric GTA 15E/776 drive systems were chosen for the tests since this size of truck represented the majority of the population of large size haulage trucks and the greatest technical development problems. Test operations have proven the feasibility of equipment designs in actual operation and preliminary economic data has convinced the operators to proceed with planning to modify their entire mining operation to use trolley assist. The initial capital investment incurred for trolley assist can be recovered in a short period of time. Following this quick capital recovery, mines will continue to realize lower operating costs through increased productivity and lower diesel consumption.

Jan 1, 1980

By George H. Roseveare

Early flotation dealt primarily with the bulk concentration of sulphide minerals or the separation of one sulphide from another and was referred to as oil flotation because oils and fats were used as reagents. The separation or concentration of one nonmetallic mineral from another developed later with the use of fatty acids. Fatty acids are organic acids present usually as glycerides in tall oil and oils of fish, peanut, cottenseed, soya beans and ethers of veg¬etable and animal origin. Some, of these oils and their derivatives were used by the Phoenicians as early as 600 B.C. in making soap, and the use of such reagents in mineral dressing is referred to as "soap flotation". The first use of oleic acid for the recovery of minerals was the early agglomeration process.' The coarse particles of ore, from 1/16 to 1/4 inch diameter, were coated with large amounts of soap or oil and after screaming from the gangue were recovered by screening, hydraulic or table classification. In 1886, a patent was issued to Mrs. Carrie Everson in which the essential feature was the constituted commingling of pulverized ore, fat or an oil, either animal, mineral, or vegetable, for the purpose of effecting .a. union of the free metal or metallic portion of the ore with such admixed material whereby the same may be changed in subsequent separation of quartz or other rock therefrom by washing or other suitable means. The patent mentions the use of red, cotton seed, caster, sperm and linseed oils, In the early use of collectors the amount of oil for froth flotation was often 2 to 6 per cent by weight of the ore rather than pounds per ton.

Jan 1, 1959

By T. M. Storer

Much has been written and many discussions held on the virtues (and evils?) of a Company-owned motor fleet. Almost invariably these dialogues take a generalized form and thereby lose something since the individual circumstances, geography and philosophy - particularly in diverse industries, is never quite the same. In considering the nature of my remarks today, I felt a departure from the general approach might be in order, and in lieu of that, will attempt to present a concise picture of how we in the Lehigh Portland Cement Company view this problem and the steps we have taken in our efforts to solve it. At the outset a word may be in order about the Lehigh Company itself, what we do, and the problems that confront us specifically as well as the cement industry at large. We are primarily a manufacturer of portland cement operating 10 plants with a total annual capacity of about 23,000,000 barrels. We market this material in 30 states, shipping both from the plants directly as well as through a system of 18 distribution terminals. About 150,000 truckloads of cement are shipped to customers each year.

Jan 1, 1969

By J. W. Butwell

Anaconda has been active In the Bingham mining district near Tooele, Utah since 1914 when the company acquired International Smelting & Refining Co. Over the years both custom ores and complex Pb-Zn-Cu ores from the district were milled and smelted at International. In 1948 Anaconda acquired the property that covers the present Carr Fork ore body. Since that time active exploration took place in the 50's followed by detailed geological evaluation in the 60's. Convinced that high grade Cu deposits would be found at depth, a deep drilling program commenced in the late 60?s, with intersection of the ore zone in 1969. Following extensive deep drilling in the early 70?s, a feasibility study was conducted in 73-74 with board approval given Sept. 6, 1974. The 9,000 t/d mine-mill complex was budgeted at $217 MM over a 5 year schedule with start-up in Sept., 1979. The following discussion includes the concentrator and surface facility portion of the budget or approximately $60 MM of the scope. Special emphasis will be made on equipment selection and layout.

Jan 1, 1980

By Kenneth V. Luza

Underclays are associated with coals found in eastern Oklahoma. These coals are of bituminous rank and are in beds of Middle and Late Pennsylvanian age. They occur in an area of about 8,000 mi2, within the southern part of the Western Interior Coal Region. The rocks of the area consist of a thick series of sandstones, shales, underclays, coals, and limestones. There are more than 40 beds of bituminous coal in the state, but only 12 of these have been mined over large areas. The 12 are, from oldest to youngest: the lower Hartshorne, upper Hartshorne, McAlester (Stigler), upper McAlester, Cavanal, Rowe, Secor, Weir-Pittsburg, Mineral, Croweburg, Iron Post, and Dawson coals. None of the underclay associated with these coal beds is known to have been exploited commercially. Nine samples of underclay from beneath the Croweburg coal were collected over a wide area in the eastern Oklahoma coal field for evaluation. The underclays are usually shades of gray with brown or orange iron-oxide staining. Although soft and plastic in some places, they are generally hard and slickensided, and contain abundant carbonized plant fragments. Bulk chemical analyses, clay mineral identification, and ceramic tests were performed on each underclay sample. The dominant clay minerals are illite and kaolinite with some mixed-layer clays. Silica (Si02) ranges from 45.96% to 59.03% and alumina (Al2O3) varies from 14.98% to 23.60%. The U.S. Bureau of Mines Tuscaloosa Re- search Center conducted the ceramic evaluations. Samples were slow fired from 1,000°C to 1,250°C in 50°C intervals. Bulk density, color, percent linear shrinkage, and percent absorption were determined for each slow fire test. All Croweburg underclay samples are best suited for building brick.

Jan 1, 1999

By N. S. Tanious

Three water-spray ventilation systems were tested to determine their effectiveness in diluting methane at a room and pillar coal mine face. These systems were: the shrouded nozzle, the spray fan, and the conventional nozzle system that is normally supplied with continuous miners. The testing was carried out in two stages. The air-moving capacities of the nozzles were first evaluated in a duct. The duct tests showed that: (1) the volume of air moved by the nozzle is controlled by the type of nozzle, the space around it, and its location relative to the duct wall; (2) the spray fan nozzles or the shrouded nozzles are both capable of moving air quantities up to 6400 cfm (3.02 m3/s); and (3) when the shroud was removed from around the shrouded nozzle, its bare nozzle moved between 9 and 65 percent more air. In the second stage of testing, a full-scale evaluation of the three systems was conducted. The three spray systems were mounted on a continuous miner which was positioned in an aboveground simulated mine entry. The simulated face tests gave the following results: (1) the standard nozzle system, which is mounted on the front end of the miner, is inadequate for diluting methane; (2) under the best testing conditions (lowest methane emission rate, highest primary ventilation airflow, and closest brattice set back), the slight improvement given by either the shrouded system or the spray fan is the same; and (3) under the worst testing conditions (highest methane emission rate, least primary ventilation airflow, and farthest brattice set back), the spray fan gave superior performance.

Jan 1, 1987

By M. F. Dibble

Borehole mining, as a viable concept for recovering deep sedimentary mineral resources, historically has been long on promise but short on results. Application of the technology seems best suited for recovery of Miocene phosphate resources on the East Coast of the US. These deposits are potentially enormous in scope and underlie sections of the coastal plain and the near offshore continental shelf. The landside portion of these deposits is generally too deep to mine using conventional open pit methods. Recovery of these minerals requires development of alternative mining systems. A significant proportion of the offshore resources appear to occur at depths below the practical limits of cutter suction dredges and may even exceed the limits of bucket ladder dredges. In most cases, these deposits are covered by overburden that must he removed and redeposited before the mineral section can be mined. The depth, thickness, composition of overburden and environmental sensitivity of a particular location would greatly impact mining system selection. It is probable that dredging would not be an environmentally acceptable mining method in many of the areas underlain by this phosphate rock. In cases such as these, new processes such as borehole mining must be evolved or the resource will remain undeveloped. A borehole mining test program, using an existing prototype borehole system, successfully extracted Miocene phosphate in northeast Florida in 1984-1985. The top of the main mineral zone was 80 m (261 ft) below the surface. Based on the results of this test program. plus subsequent laboratory testing and engineering studies, a new prototype borehole mining system was designed in 1986. This design incorporated several innovations of offering potentially higher productivity and lower power consumption than was achieved with the previous test unit. Economic restraints at that tine prevented the actual construction and testing of this new design.

Jan 1, 1992

By Immo H. Redeker

Introduction In October of 1961 Texas Gulf Sulphur Company contacted the Asheville Minerals Research Laboratory, of the North Carolina State University, and requested assistance in the development of a phosphate ore body in Beaufort County. The main: objectives of the work at Asheville were defined as follows. The Laboratory should try to determine on the basis of drill core samples, submitted by Texas Gulf Sulphur Company, the highest grade of phosphate concentrate that could be obtained with high recoveries from North Carolina ore, the thickness of the ore body that could be treated economically in one section, and the quantity of concentrate that could be obtained per acre. The results should be compared with data obtained from the treatment of Florida ores. It was important to use mineral preparation methods during the initial laboratory work that could be verified later by continuous pilot plant operation and could be scaled-up for a commercial plant. The writer was in charge of the project at the Asheville Minerals Research Laboratory and participated in the development of one of North Carolina's greatest mineral resources from the first drill core examination to the production of the first poundage and tonnage quantities of flotation concentrates. The Lee Creek phosphate mine and mill was put in operation in early 1966 and can produce three million tons of phosphate concentrate a year. Sulfuric- acid, and phosphoric acid plant, diamonium phosphate plant and triple super phosphate plants came on stream in 1967; at that time they were the world's largest. Now that a phosphate industry has been established in the State, it]

Jan 1, 1971

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By R. R. Rollins

Field data combined with theoretical derivations were used to design a surface mining operation using explosive casting. Calculated casting results varied from 44 to 65% depending on bench height. This model was then applied to field tests at an eastern, contour, coal stripping operation. Results from five experimental casting shots are summarized and compared to conventional mining methods. The overburden material removed by casting ranged from 35 to 65% with an average of 50%. Fragmentation was improved and it was demonstrated that this procedure is economically attractive with a resultant cost savings of $.59/m3 or $5.93/t of clean coal.

Jan 1, 1987

By M. K. Larson, J. K. Whyatt

Proper employment of numerical stress analysis design tools is based on the demonstrated ability of a model to capture key elements of the geologic site model and accurately simulate how these elements interact with a mine design. While these tools have progressed markedly, they are, at heart, a gross simplification of the abundant complexity of a natural setting and its response to mining. A generic deep longwall site model was developed that includes aspects of the geology of deep coal mines in the Wasatch Plateau and Book Cliffs coal fields of Utah. The site model contains a set of common features and observations of how these features typically respond to mining. This site model was the basis for evaluating use of empirical, boundary element and volume element stress analysis tools to analyze the distribution of stress around a deep longwall panel. More specifically, this evaluation examined shifting of stress to panel abutments and gob, distribution of stress in the abutment and deformation and failure of bridging strata. Measurements of abutment stress changes at two sites in the Wasatch Plateau region were used to illustrate model calibration. Overall, these comparisons highlight the considerable differences between methods. Volume element tools can incorporate considerable detail and have fewer underlying assumptions, but this detail carries a considerable computational cost. Boundary element tools are much more efficient. However, this efficiency comes at the cost of added assumptions. These assumptions were challenged by the presence of a strong sandstone unit in the overburden, leading to boundary element results that depart significantly from volume element results. Empirical rules are the simplest, but are even more burdened by assumptions, many of which are implicit in underlying cases. Insight into the nature and impact of underlying assumptions in each method is essential to the proper use of results in mine design.

Jan 1, 2009

By W. D. Monaghan

One of the major ignition sources of fires and explosions in underground U.S. coal mines is flame cutting and welding. On May 20, 2006, a flame-cutting operation at the Darby Mine No. 1 led to an explosion that resulted in five fatalities. On Jan. 22, 2003, another explosion caused by flame cutting at the McElroy Mine resulted in three fatalities and three serious injuries. On June 22, 1999, a flame-cutting and welding operation at the Loveridge Mine caused a fire that required sealing the mine for an extended period. On March 19, 1992, an explosion initiated by welding at the Blacksville No. 1 Mine resulted in four fatalities. These examples of mine incidents demonstrate the need to conduct research to develop best practices for safely conducting flame cutting and welding in underground coal mines. The National Institute for Occupational Safety and Health (NIOSH) conducted a study on fires and explosions in underground U.S. coal mines that were caused by flame-cutting and welding operations to determine the root causes of these types of incidents. The methodology included interviewing mining personnel who perform flame-cutting or welding operations in underground U.S. coal mines and visiting mines to observe these operations. In addition, MSHA reports of investigations and accidents statistics were analyzed. The findings were used to identify and compare differences between flame-cutting and welding practices and techniques in small and large mines, eastern and western mines, low-seam and high-seam mines, room-and-pillar and longwall mines, and between experienced miners and new miners. As a result of this study, best practices were developed to reduce the number of ?res and explosions caused by flame cutting and welding. This paper provides a summary of this research.

Jun 1, 2010

By Howard M. Wells

Portfolio theory and the Capital Asset Pricing Model (Modigliani and Pogue, 1974) provides an investment decision rationale which takes account of the benefits of diversification in reducing effective risk, but this theory does not consider the financial capability of the firm to survive large losses if they should occur. The LL-curve theory (Wells, 1976) takes the latter consideration into account, but it does not provide for the effects of diversification. This paper pro- poses a general theory to account for both diversification and the financial capacity of the firm to underwrite the investment risks.

Jan 1, 1983

By B. A. Ashley

Rosebud Mining Co. (RMC) is based in Kittanning, PA. As of October 2013, it operated 24 active underground mines in Pennsylvania and Ohio. RMC?s Tracy Lynne Mine is located approximately 5 km (3 miles) east of Apollo, PA, (approximately 48 km or 30 miles northeast of Pittsburgh, PA) and mines the Lower Kittanning coal seam. The coal seam averages approximately 1.1 m (45 in.) thick and is of steam coal market quality. The mine is a single unit operation. Coal is mined using a continuous mining machine and is transported from the face to the conveyor belt using a continuous haulage system. Typical ventilation in the mine consists of an exhausting fan that draws air through isolated intake, belt and return air courses (Fig. 1).

By R. Casagrande, L. R. Moore, J. Sessoms, P. Macy

"A South American nickel mine has been producing a nickel concentrate that also contains copper and iron (15 percent nickel, 5 percent copper, 28 percent iron) for the end use of stainless steel production. A critical specification for the quality of the steel is the iron to magnesium oxide ratio (Fe:MgO), with a ratio of below 3.5:1 having a substantial impact on the nickel value. This nickel concentrate producer historically mined the sulfur-rich mineral pyroxenite but had been forced to move to a magnesium oxide/silicate (MgO/SiO4)-rich mineral harzburgite, resulting in nickel recovery being reduced to around 50 percent to maintain the required Fe:MgO ratio. In this study, ore mineralogy, water chemistry and the chemistries associated with the flotation process were evaluated in an effort to increase nickel recovery while maintaining the Fe:MgO ratio according to specification. New chemistry was introduced with the specific goal of complexing with the MgO and retarding its interaction with the bubble flow in the flotation process.IntroductionGlobal nickel production in 2011 totalled to 1,770,000 tons, according to the Raw Materials Group (2013). From 1950 to 2005 there was an aggregate 4.4 percent increase in nickel production, with a market-value high of $1,000/ton nickel reached in 1998 (Mudd, 2010). Approximately 65 percent of nickel production is consumed by the stainless steel industry, with 12 percent of the remainder used in super alloys or nonferrous alloys. Nickel imparts corrosion resistance, temperature resistance and other benefits to its alloys (Table 1) (Oberg, 1996; Cáceres, 2003). The dominant types of nickel-bearing ores are classified as either laterites (limonite: nickel-iron oxide; and garnierite: nickel hydrosilicate) or magmatic sulfide deposits (pentlandite: iron-nickel sulfide) (U.S. Geological Survey, 2013a, 2013b). In 2012, about 40 percent of the nickel ore mined was sulfidic-type ores and about 60 percent was laterite. The processing of laterite ore is expected to continue to increase as the availability of sulfidic ores continues to decline."

Jan 1, 2015

By Arash Dahi Taleghani, Pooyan Asadollahi

Slurry walls, also known as diaphragm walls, are commonly used as temporary and/ or permanent support of excavation for construction of shafts in soft ground. Through case history, this paper demonstrates economical benefits of disregarding long-term structural contribution of the slurry walls particularly in severe environmental condition. In addition, for temporary slurry wall shaft systems, possibility of using steel fiber, synthetic fiber or additives in lieu of temperature/shrinkage reinforcement is discussed with detailed analysis. Finally, the paper will be concluded by introducing unreinforced slurry wall shaft system using newly developed/patented concrete additive.

Jun 13, 2021

By Rusdian F. Kayadoe, Michael B. Hutahaean, Sandi A. Firmanulhaq

Deep Mill Level Zone (DMLZ) mine is located in Papua, Indonesia operated by PT Freeport Indonesia (PTFI) with approximately 1,500m below surface. The ore body extend 1230m width from northwest to southeast by 300–400m length east-west. The mine currently produces at full production of 80,000 tons per day until 2041. Part of the extension of current active production block and mitigation sits 225 m above the production block 1 (PB1) and 2 (PB2A and PB2B) which located in 2590L, named production block#4 (PB4W and PB4E). As the mining continues at PB1N, PB2A, and PB2B, the cave growth vertically and laterally to south perimeter of the footprint affecting the integrity of PB4W/E footprint. Currently, PB1N and PB2N are in active production whilst PB4E development continues from South against PB1N cave, as can be seen in Figure 1. A slight overlapping between PB1E and PB4E intentionally to promote cave connections between the two caves. Despite low stress model in the PB4E near existing caves, the PB4E cave propagation expected by the large strain and will be propagating in tension failure to the PB1E cave. Hence, designing the drill drives edge and first drawbell location as close as possible to the PB1E cave is necessary to guarantee cave connection. Since PB4 footprint development commenced in December 2021, there have been a substantial cave propagation observed from existing downhole monitoring. As the PB4 footprint designed perpendicular to existing DMLZ PB1 cave, adjustment of the development edge design is necessary to reflect the progression the loosening zone propagation to the South. Final life of mine design for extraction, undercut and ventilation level were decided in October 2023 considering latest update cave (fractured/ yielded) boundary. Numerical model of potential rock mass damage also aids the decision to locate the drift edge design. Predicting the cave progression impact is critical to both production at the existing active blocks and development of the new area. The use of cave shape data from year 2021–2023 gives an insight of the cave shape estimation at the end of 2024 when the undercut drift will be developed. This paper presents actual cave growth of active production blocks and its relative vertical progression impact as the result of drawing in order to plan for acceptable rate and provide mitigation once the minimum acceptable rock mass condition and minimum pillar threshold achieved.

Feb 1, 2025

By T. R. Kostecki, A. J. S. Spearing, G. S. Esterhuizen, K. V. Jessu

Discontinuities are geologic occurrences in rock and when present within a pillar, reduce the strength of the pillar. Empirical formulas that are commonly used to determine pillar strength do not explicitly take into account the presence of discontinuities and thus can overestimate the pillar strength. The effect of discontinuities on the strength of pillars has been investigated using numerical models, but in these models, the discontinuity strike was parallel with the pillar faces. In this study, fully three-dimensional hard rock pillars were simulated using numerical modeling to understand the effect of the discontinuity dip direction on square and rectangular hard rock pillars. Based on the results, recommendations to assess a pillar’s strength in the presence of a discontinuity are discussed.

Jan 1, 2018

By Jin Zhao, Weijie Wei, Jinwang Zhang, Xiaohang Wan, Shengli Yang

Reasonable mining-retention ratio (ratio of room width to pillar width) is vital to recovery rate and stope stability in underground mining of solid potassium salt deposits. Based on the engineering background of a solid underground potassium salt mine in Khammouan Province, characteristics of the plastic zone of the pillar, and the variation of recovery rate and stope stability under different mining-retention ratio conditions are discussed. The results show that compared with current room width (8 m) and pillar width (10 m), the optimization method of improvement of stope stability. The research results are helpful to recover as much solid potash resources as possible under the premise of ensuring the stability of stope.

Jun 25, 2024