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By D. G. Feuerbacher

Maximizing production of oil from sandstone petroleum reservoirs is aided by use of non-damaging drilling and workover/completion fluids. Various tests, such as X-ray diffraction, scanning electron microsccpy, petrographic microscopy, porosity and permeability are conducted on core samples. Several fluids are selected, on the basis of these results, for return permeability tests. A report is then written giving the results of all tests, potential reservoir problems, and recommended practices for exploiting the reservoir.

Jan 1, 1983

By M. Szatkowski

The studies conducted by Laplante and coworkers (1983,1984) with flotation of galena and quartz, showed that the rate of flotation reaches a maximum value for a certain intermediate value of the air flow rate. A similar, though less pronounced, effect of the air flow rate on the selectivity was also observed. Luttrell and coworkers (1985) showed that the size of air bubbles has a significant effect on both rate and selectivity of coal flotation. These two studies used flotation tests in which no froth layer was formed. This note describes effects of the air flow rate and bubble size on the rate constant and selectivity of coal flotation, using a mechanical laboratory machine with a froth layer allowed to form. Experimental A raw, dry ground Illinois coal was used in flotation tests. The size distribution and ash content of the samples are : Particle size (µm) Weight (%) Ash (%) +600 4.30 41.7 600/200 43.05 19.0 200/74 22.25 16.0 - 74 30.40 24.0 average 20.8 No. 2 fuel oil with a concentration of 200 g/t of feed and M.IBC with concentrations of 11, 22, and 33 mg/L of pulp (128, 256, 384 g/t, respectively) were used as flotation reagents. Tap water with a natural pH of 7.0 to 7.8 was used throughout the tests. No other reagents were used. The starting concentration of coal in the pulp was 86 g/L. A standard flotation machine with the pulp volume of 5 L was used in experiments. Three different air flow rates were employed, resulting in three air escape velocity values of 14, 30, and 70 cm/min. The air escape velocity (AEV) was calculated as the ratio of air flow rate (Qa) to the slurry surface (Ss) in the machine. For a comparison, the upper limit of air escape velocity values used in commercial size flotation machines is between 70 and 100 cm/min (WEMCO, 1983). A machine consisting of a generator of fine bubbles and a separatory column, described elsewhere (Szatkowski and Freyberger, 1985), was used for some tests where bubble size lower than that generated in the standard machines was required. Prior to a flotation test, a batch of coal was mixed with water in the flotation cell for at least 10 minutes to ensure that all coal was thoroughly wet. Then the specified amount of fuel oil was added and the pulp was conditioned for five minutes. The specified amount of MIBC was added 15 seconds before the end of conditioning. After air flow was opened, flotation products were collected for 0 to 30, 30 to 60, and 60 to 120 second intervals. Content of the combustible material in samples was determined by subtracting the ash content from 100%. A photographic technique was used to determine initial distributions of bubble size generated by the flotation machines. Bubbles generated in water containing the prescribed amount of the frother but no solids were photographed against a black background. The average diameter and standard deviation were then calculated for each set of flotation conditions. The method gives results reproducible within k 5% of the average bubble size. Table 1 contains information about the measured bubble size distributions expressed by the average bubble diameter (Db) and the standard deviation ( SD ) . Results and discussion The general formula for calculating the flotation selectivity is: o = recovery of coal/recovery of ash (1)

Jan 1, 1988

By Trevor R. Ellis

In this paper a method is presented for estimating the specifications of sandstone-type uranium deposits which will be the minimum that can meet a company's economic requirements for investment. The method interpolates and extrapolates from available cost data for typical mines to provide the costs for a wide range of possible deposits. The range of deposits achieving the company's required discounted cash flow rate of return on investment at an estimated uranium oxide selling price is then calculated, thus giving the complete range of deposits which meet the minimum required specifications. Prospects are then evaluated by comparison with these minimum requirements.

Jan 1, 1979

By Frank A. Aplan, Barbara J. Arnold

INTRODUCTION Coal flotation is a fine cleaning process usually present in all but the older preparation plants. It is typically used to clean either the -28 mesh (-500 pm) or - 100 mesh (- 150 pm) raw coal fines, and it is the only commonly used preparation method that can effectively recover coal particles finer than, say, -200 mesh (74 pm). To place coal flotation into its proper context, it should be noted that the bulk of the raw coal sent to preparation plants is treated by gravity concentration. On a worldwide basis, from 0 to 40% of the raw feed to a preparation plant is treated by flotation, while in the US only about 10% of the raw coal feed is floated. Roughly, 50% of US preparation plants use flotation to process at least a portion of the raw coal feed, although most of the newer plants incorporate flotation into their flowsheet. Historically, the process has only been used to treat the raw coal fines naturally occurring after mining and a minimal amount of crushing at the preparation plant. This is in sharp contrast to ore flotation where, typically, the entire tonnage is first ground and then floated. As a consequence, of the approximately 500 Mt of raw coal treated by preparation in the US yearly, it is estimated that only about 40 Mt are sent to the flotation circuits (producing -30 Mt of clean product).'O1 This small tonnage compares with about 400 Mt of ores treated annually in the US by flotation. By analogy to ore flotation, where it is an effective, large tonnage, low cost, easy to operate process, its potential in coal preparation should be great. Reasons why it has not achieved a greater potential in coal preparation have been attributed by Aplan 102 to: 1. Indifferent results in removing sulfur and ash. 2. Difficulty in floating low rank and oxidized coals. 3. Inability to make clean separations with the finer sizes. 4.Inability to clean slurries with a high percentage of clay. 5. Problems with water chemistry. 6. Competition with effective, gravity coarse-cleaning methods. 7. Low value of coal historically. 8. Dewatering problems. 9. Transportation and storage problems with coal fines. 10. Marketing problems with coal fines. 11. Inadequate attention to the process in the plant. 12. Inefficient use of capital and operating monies. Inadequate attention to procedures, machines, and processes.

Jan 1, 1991

By Richard Keith Corbin

INTRODUCTION The purpose of this paper is to assist explorationists and entrepreneurs in deciding whether to invest in gold mining concessions in lesser developed countries (LDCs) and, if so, in deciding which LDCs provide the best opportunity. The first decision is difficult in the face of common notions that foreigners are not welcome - only their money; that the best gold deposits were worked out long ago; that even if you survive the mal- aria, and in spite of 51% control by a national partner, you manage to profit from a new discovery, they will tax your operation to death, or worse yet, nationalize it. Close scrutiny of what happened to foreign investors in LDCs during the “Mexicanizationt” of the 1930's through the "Zambianization" of the 1970's shows that many of these concerns are well-founded historically. But the trend of knowledgeable writers over the past decade seems to be away from this gloomy foreboding, and towards a cautious optimism, based on recently perceptible metamorphoses in certain LDCs' conditions and attitudes. If the explorationist or entrepreneur decides to seize on this apparent opportunity, the second decision -- which of these certain LDCs to invest in -- is doubly difficult, because of the formidable tasks of coming up with a rating system to provide rational comparison, and then gathering enough current, reliable information to apply. To assist in these tasks, a suggested rating system and case studies, respectively, follow a brief review of recent writings on the subject

Jan 1, 1987

By W. J. Schlitt

Introduction A major uncertainty in assessing the potential for heap-and dump-leach projects is how to determine the extraction-rate curve for the recovery of the mineral values from coarse ore. Such material could either be run-of-mine (ROM) or primary crushed ore. The problem with field testing coarse ore, especially for new projects, is the large scale and extended leach times needed to accurately determine the final extraction-rate curve. At least 5 x 103 to 5 x 104 t of representative ROM ore are typically required for a copper test heap, and much more is often used. Kennecott, for example, recently constructed a 0.9 Mt (1 million st) ROM test heap at the Bingham Canyon Mine in Utah. In such coarse ore operations, the ultimate level of extraction will require a leach cycle that can extend from several months to a few years. Quite often, project development schedules do not provide the luxury of mining such large quantities of material or running such long tests. Instead, test data are usually limited to results from column leach studies on relatively fine ore, often with a top size that does not exceed 25 mm. Maximum leach times are also short, typically less than a year before an initial decision is needed on project viability. Proposed method One approach to estimating the recovery from a coarse ore leach is to assume that the leach solution will have some ultimate penetration distance into the rock. Then, the final level of mineral extraction in this "leached rim" will equal the ultimate level of extraction identified in various testing programs. Obviously, if the radius of a given rock fragment is less than the penetration distance, that fragment will be fully leached at the end of the operation. With larger rock sizes, the percent recovery will fall off as the size increases and the fraction of unpenetrated rock mass increases. Such an approach sounds simple but is likely to be complex when applied to a real project. For example, the penetration distance will be a function of both the rock characteristics and the effective length of the leach cycle. The important rock characteristics include rock porosity, the degree of internal fracturing and the mode of mineral occurrence. With regard to the latter, penetration is likely to be greater if the leachable mineralization occurs on fracture surfaces or in veinlets, as opposed to fine grains uniformly disseminated throughout the rock mass. An estimate of penetration distance may be derived from column or heap tests by noting the depth of solution penetration into the larger rock fragments after three, six and 12 months of leaching. While the penetration rate is ore specific, something on the order of 10 to 20 mm/y may be appropriate for competent, primary copper (chalcopyrite) ore. For gold in tight quartz, the rate may be about the same. Copper oxide ores and gold that is hosted in a more porous rock matrix are likely to have penetration rates that are at least two to three times higher, and an even higher rate should be appropriate for uranium hosted in sandstone. As noted above, the length of the effective leach cycle is likely to be measured in years. On this basis, the ultimate penetration distance (dp) would vary from less than 50 to several 100 mm when a particular ore is leached to exhaustion. Several sets of mathematical manipulations are necessary to convert a rock size distribution and corresponding value of dp into an estimated extraction-rate curve. The first step is to break the ROM size distribution down into intervals and then calculate the radius for the mean rock size in each interval. This is shown in Table 1 for rock sizes up to 1.75 m (about 6 ft) in diameter. The next step is to calculate the volume of unleached core and the fraction of rock that is leached. This is done for the following three values of dp: 25, 100 and 250 mm. Results are shown in Table 2. The third step is to select the ultimate level of recovery that will be achieved in the fraction of material that is effectively leached, i.e., the outer zone that is penetrated by the leach solution. This is clearly a site-specific factor that can only come from metallurgical test results on representative ore

Jan 1, 1998

By B. M. Davis

The problem of assessing uncertainty in grade estimates has been a topic of research and debate since kriging was first introduced. Use of the kriging variance was originally touted as all that was necessary in order to produce confidence limits for grade estimates. Since then the prudence of using the kriging variance for confidence interval estimation has been questioned, and techniques such as the jackknife and conditional simulation have been proposed to develop "true" interval estimates. This paper compares the differences between interval estimates produced by the three methods and makes conclusions about the performance of confidence intervals produced by the different methods.

Jan 1, 1997

By D. W. McGlashan

[Becaune ad ;the extennive nwrgace anea od u.X chopRex2 in emmiom, emuhion dRokdtiun oddeh poaaibWa don ;the necove/ry ul; BinQey divided minmcd pa~LL&a, Swr~aC;tavu% mu2 akab&ze ;the emu-bion and poan an an ad~i&y don ;the dahed mina&. OZhm hpohtant acto oh me ;the naAwre od ;the em&iun inghedied, ;the nQRaLLve volwna 06 the colz;tinuoun and dincon;tinuoun phaa, ;the change aign bonne by the dnuplm ad the o.X phae, ;the z&a pu;tenticd od ;the dahed minmd, the ~olubRe am curdent 06 khe wcr;ten-pulped and gnuund onu. AX on nem ;the AoQRec;trtic point on ZPC, ;the doubRe Raym about ;the minW-&den oil dnopR& i~ neduced in ;tkichnaa and the dn_opR& wU codace and continue to aggnegate into agglommaka . Invetmion dnom O/U ko W/O can UCCWL depending upon a nwnbm a 6 in;tmQeaked daotuh~. EReothokin&Lc inn&wnenta;t/iun to meanwre ;the ZPC, QReckProphonQkic mobilXty z&a pokeW and hakerning cuttrtent me invcduabte in ;the appficatiun ud emu-bion dRo;tat;iun ;to m&~c and nonm&~c minu.&. ] INTRODUCTION Recent interest in emulsion flotation can be ascribed to the necessity of selective separation and concentration of extremely fine minerals. That which follows is an elucidation of the many factors-in emulsion technology in application. In applied surface chemistry for solid-particulate separation and concentration, emulsion flotation is one of the intermediate processes between froth flotation, in which the continuous phase is water (99.5 to 99.999% by volume), and particulate extraction, in which the continuous phase is almost entirely an organic-fluid. Another intermediate process is particulate exchange (or partition) in which immiscible or very limitedly miscible phases are in controll¬ed but appreciable volumes. Surfactants are common to all of these

Jan 1, 1968

By C. Manea

During exploration and exploitation of complex ores, waste dumps that contaminate severely the environment and in particular soils, are genetated. The most hazardous pollutants are heavy metals and metalloids (e.g As, Mo, Pb and Cu) as well as radionuclides (e.g U and Ra). Ecological remediation aims, among others, at decontamination of soils by preserving important properties as soil fertility.

Jan 1, 2008

By Zhou Aimin, Xie Benxian, Song Chongkai, He Zhexiang, Zhang Changqing

Concrete pump is commonly used to transport paste and high-density backfill over relatively long horizontal distances. However, since the investment in concrete-pump and its operating cost are quite high, the application of paste backfill is limited in China. For seeking an alternative paste backfill technology, squeezed-transport technique and equipment has been developed. The new technology has been adopted at Tonglushan Copper Mine of Daye Nonferrous Metals Corp for overcoming a difficulty of long-distance transportation of high-density cemented tailings fill. It proved to be an alternative paste backfill method with simple backfill system, saving investment and low operating costs.

Jan 1, 2001

By Scott Burchmore, Jason Emery, Michael Melville

This paper presents best practice data management for monitoring of extensometers in Australian underground coal mines. This work is based on the combined experience of the authors and the available literature. The information and ideas presented herein are a culmination of many years of database development founded from mine site-based ground support design, underground inspections, extensometer data management and data analysis. Each author has more than ten years’ experience as a practicing geotechnical engineer at several operating industry wide research project focused on utilizing mechanical extensometer data for data analytics and machine learning. Principal hazard ground control management plans, stipulated in Australian coal mining legislation, outline the mandatory installation of ground monitoring devices during routine mining activities. After 20+ years of data gathering the barrier to implementing modern data analytics remains the challenges imposed by inconsistent data structure and data quality. This inconsistency is still routinely found between sites, companies and mining regions. Through the relatively simple application of a standardized database structure, the potential to shift from simplistic tracking of trigger levels, siloed at individual sites, to an industry standard method for compilation & comparison of monitoring data is enabled.

Jun 25, 2024

By E. J. Wasp

The transportation of mine products incurs a cost which can have a significant bearing on the competitive position of the manufactured product. Clearly, the engineer who is responsible for selecting a mode of transport must base his decision, at least in large measure, on economic comparisons. And today, the engineer usually finds that he has to choose between several modes, including slurry pipelines. For the purpose of an initial economic screening, it is not necessary to go deeply into the technical details of a slurry system. What is usually needed, however, is a method with which the engineer can make a quick economic assessment to determine whether a slurry pipeline should be included in the detailed final analysis. This paper presents a rapid method for making such a preliminary appraisal. Experience with commercial slurry pipeline systems in several countries and with a variety of solid materials has shown that if several assumptions pertaining to critical design variables are made, it is possible to roughly estimate the transportation cost of a slurry system on the basis of only three factors: the specific gravity of the solid material, the length of the system, and the quantity of slurry to be pumped each year. It is assumed that conditions pertaining to field construction, corrosion inhibition, and the like, will be favorable. This assumption is critical because any factor which analysis shows to be particularly unfavorable will raise the total transportation cost estimate which this paper establishes. If the engineer employs this estimating procedure and decides that the use of a slurry pipeline is economically unattractive, his chances of being wrong are small. Conversely, if his quick appraisal indicates that the use of a slurry system is economically favorable, the engineer is justified in committing the time and money needed to make a more detailed investigation of the slurry pipeline potential.

Jan 1, 1969

By Marco Sonnenschein

Belt conveyors for tunnelling are industry standard and key to high performance of TBM drives. Conveyors reduce traffic in the tunnel and increase safety in the tunnelling process. This paper will illustrate the state of the art of tunnel belt applications in very long and curvy tunnels like the Coxwell Bypass Tunnel Project in Toronto, Canada.

Jun 23, 2024

By Mustafa Kumral, Samer Hmoud

Effective management of ore loss and dilution is essential for successful grade control and short-term mine planning due to their significant impact on the economic, environmental and technical aspects of openpit mining operations. Factors influencing ore loss and dilution fall into two categories: (1) controllable factors, like mine equipment selectivity and blast design, and (2) uncontrollable factors, such as spatial heterogeneity of ore and waste blocks on a bench. This paper focuses on the second category by applying spatial entropy concept to capture heterogeneity at the scale of selective mining units. In this paper, global spatial entropy is used to assess the impact of spatial heterogeneity between ore and waste blocks on the magnitude of ore loss and dilution, while the local spatial entropy can guide the allocation of blast movement monitoring balls preblast. High values of the global spatial entropy indicate increasing potential of ore loss and dilution, which reduce profit. Furthermore, through a number of case studies, the study investigates the relationship between spatial entropy, cutoff grades, blast movement, dig-limits optimization model running time, and profit. The results show that changes in cutoff grade and blast movement can significantly affect spatial entropy post-blasting and increase ore loss, dilution and profit reduction, revealing the need for controlled blasting at specific bench sections. Additionally, the results demonstrate an exponential increase in profit reduction due to ore loss and dilution with a rising global spatial entropy

Jun 1, 2024

By Wolfgang R. Wawersik, Charles M. Stone

INTRODUCTION Hydraulic fracturing was applied in horizontal drill- holes in the Salado salt formation near Carlsbad, New Mexico. Tests were performed approximately 650 m be- low surface to validate the design of a Waste Isolation Pilot Plant (WIPP) for the disposal of radioactive waste from defense activities of the United States. Hydraulic fracturing was performed primarily to determine whether the virgin in situ stress state at the WIPP site is isotropic and whether the magnitude of the virgin in situ stresses corresponds to the weight of the overburden. Beyond these limited objectives, however, measurements are now being reviewed to evaluate the usefulness of hydraulic fracturing in salt formations in general. Such measurements are desirable to determine stresses induced by mining and to monitor time-dependent stress changes around underground excavations in salt masses. Hydraulic fracturing was chosen because it is the only technique that can be readily applied in virgin ground far from existing mine openings. The method also appeared to yield directly the magnitude of the least compressive principal stress, regardless of the mechanical properties of the rock mass. A knowledge of the least principal compression was deemed sufficient to establish the isotropic character of the virgin in situ stress field under two conditions: (1) the Salado formation was mechanically isotropic, i.e., the fracture toughness did not vary normal and parallel to bedding so as to permit preferred fracture paths, and (2) pressure measurements were combined with determinations of the orientations of hydraulic fractures and, more generally, with observations of the characteristic hydraulic fracture patterns. The latter was possible by locating the drillholes for hydraulic fracturing along the axes of drifts that were mined several months later. The success of general stress measurements by hy- draulic fracturing of salt subjected to anisotropic stress conditions was and remains less certain. While the smallest in situ compressive stresses should still be equal to instantaneous shut-in pressures, there may not exist any unique relationships between the greatest and intermediate in situ principal stresses on one hand and the primary breakdown and fracture reopening pressures on the other. This problem would be a consequence of stress relaxation due to salt creep around any drillhole in which hydraulic fracturing tests are conducted. Moreover, lab- oratory measurements suggest that creep and stress relaxation could be so rapid that it might not be possible to interpret the measurements for anisotropic stress conditions even if hydraulic fracturing were performed as soon after drilling as possible. The primary objectives of the stress measurements in the WIPP were met by combining hydraulic fracturing tests with finite element analyses of the effects of salt creep, observations of hydraulically induced fracture pat- terns, and laboratory hydraulic fracturing tests on oriented core. In situ hydraulic fracturing tests in long drillholes were preceded by several trial tests to establish the best test methods and to demonstrate that salt could be fractured. The resolution of anisotropic stresses and special problems, e.g., in the identification of instantaneous shut-in pressures, are currently being addressed in hydraulic fracturing tests in a mine pillar, in tests following different delays between drilling and fracturing, and in tests with different frac fluids. This paper will review the main results of the analyses above and of one group of hydraulic fracturing tests at different distances from existing excavations. The paper will conclude with a brief description of the geometry of the hydraulically induced fracture patterns. Because of space limitations, it will be assumed that the reader is familiar with hydraulic fracturing or will first consult other sources (e.g., Cleary, 1980; Gronseth and Detournay, 1979; Zoback and Haimson, 1983; Nemat-Nasser et al., 1983; Kim et al., 1984)

Jan 1, 1986

By Chas. F. Jackson

INTRODUCTION This paper is one of a series dealing with mining methods and costs, prepared and published under the sponsorship of the US Bureau of Mines (USBM) and made possible through the cooperation of many mining companies in the US and foreign countries. The data presented have been obtained from papers previously published by the Bureau dealing with indi¬vidual mines, from the technical press, and from infor¬mation furnished by the mining companies. DEFINITIONS Top Slicing The term top slicing is commonly applied to that method of mining in which the ore is extracted by a series of timbered slices, beginning at the capping; the slice is caved, bringing down the capping to the floor of the slice; the succeeding lower slices are mined up to the overlying mat or gob, consisting of an accumulation of timber from the upper slices and broken capping, which follows the mining progressively downward (Fig. 1). The term has also come to be applied to the mining of ore bodies only one slice thick, by a series of tim¬bered slices one beside the other. The timbers are blasted out, bringing the capping down to the floor, suc¬ceeding slices being driven alongside those already caved (Fig. 2). In USBM classifications of mining methods, top slic¬ing has been placed under the heading caved stopes. The method differs fundamentally, however, from the cav¬ing methods of mining in which the breaking of the ore itself is accomplished by caving action induced by undercutting. In top slicing the ore is broken by the ordinary methods of drilling and blasting, and only the capping is broken up by caving. Main Levels or Haulage Levels Main or haulage levels are the main extraction levels on which the ore is hauled to the shaft. These levels and the shaft constitute the preliminary development of the mine.

Jan 1, 1982

By S.K. SHARMA, ANKUSH GALAV, G.S.P. SINGH

This paper proposes a hydromechanically coupled numerical modeling approach to quantify the influence of various parameters on the mechanical and hydraulic performance of protective water barrier pillars (PWBP). The results are validated for two case studies having pillar widths of 16 and 42 m at cover depths of 136 and 190 m and water heads of 25 and 141 m. The outcome can be significantly helpful in the rational design of these pillars in underground coal mines.

Mar 1, 2023

By Dennis Arthur Fenton

The hydrodynamic bearing was first described by, Reynolds in 1886, and has since been widely applied in motors, engines, compressors, hydroelectric turbines, and other machinery. For various reasons, the hydrostatic bearing currently predominates in the large, horizontal rotating machinery field (e.g. grinding mills). However, the current industry trend towards increasing size and energy efficiency has resulted in renewed interest in the advantages of hydrodynamic bearings. These advantages include structural improvements, reduced bearing friction, improved bearing life, improved transient characteristics, and reduced lubricant pump size and power requirements. In particular, the higher energy efficiency results in substantial savings on large machines. This paper describes some aspects of the design of large hydrodynamic bearings based on thirty years, of experience in the design of such bearings on grinding mills ranging in journal size from 2.1 m to 6.4 m, see Table 1.

Jan 1, 1985

By MOHAMED WAEL, ANASTASIA GOEVA, ABDULAZIZ TAWFEEQI, IHTISHAM TARIQ, S. M. ZAKIR HOSSAIN, NIAZ ALI KHAN, MUHAMMAD FAISAL IRFAN

Carbon dioxide (CO2) mineral carbonation is a carbon capture and storage (CCS) technique that can sequester CO2 permanently in the form of durable carbonate minerals. Extendable permanent storage capacity — onshore and offshore storage reservoirs — and the exothermic nature of the mineralization process make this technique attractive [1]. In the present study, maximum carbonation efficiency for aqueous mineral carbonation was achieved through two steps: optimization of calcium leaching from cement kiln dust followed by the reaction of pure CO2 with the calcium hydroxide precipitates formed by hydroxylation using sodium hydroxide (NaOH). Optimization of calcium leaching is highly important, as it is a rate-limiting reaction step in mineral carbonation and also influences and enhances the carbonation efficiency. To maximize the calcium leaching as well as carbonation efficiency, response surface methodology — with either Box-Behnken design or central composite design — was applied. It is the most well-known statistical, fully randomized and bias-free method for determining optimized conditions using quadratic regression models.

By Wedam Nyaaba, Edward Wellman, Angelina Anani, Eduardo Cordova, Sefiu O, Cluber Rojas, Adewuyi

This study aims to investigate the optimal method for ore pass construction in incompetent rock masses. We evaluated the conventional and raise boring methods based on safety, efficiency, excavation control and ground support for ore pass construction. A stability analysis using analytical Q-raise (QR method) and kinematic analysis methods for ore pass construction with a raise borer before and after grout injection of the rock mass was performed. As a case study, an ore pass (diameter of 3 m, depth of 100 m) within an incompetent rock mass was considered to gain further insight. The rock mass was characterized according to the classification methods Q Barton, rock quality designation (RQD), rock mass rating (RMR) and geological strength index (GSI). The grout intensity number (GIN) method of grout injection was used. The safety factor of less than 1.075 obtained before injection was lower than the acceptance criteria. However, grout injection before raise borer excavation resulted in a safety factor greater than 1.5. Using a raise borer without pre-grouting indicated that the maximum unsupported diameter (MUSD) of the ore pass was less than the required 3 m and equal to or larger than 3 m post-grouting.

Dec 1, 2024