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By R Lastra, J. A. Finch

An artificial standard material has been developed and used to examine the robustness of some of the procedures for correcting stereological bias. Different mineral liberation distributions were constructed from the standard material. The particles were mounted, sectioned and the sectioning data measured with an electron microscope interfaced to an image analyser. Four correction procedures were used and the corrected and true liberation distributions compared. It was found that the large-sections correction provided a simple, uncorrupted correction. Hill's fast approximation performed well except in cases of narrow liberation distributions; the sectioning data appeared to support the assumption that the locked section and locked particle distributions are identical. Barbery's correction performed well in most cases, having difficulty only with the very irregular liberation distribution. The PARGEN correction was able to provide a good estimate of the true amount of free material, but had difficulty in predicting the locked particle distribution

Jun 21, 1905

By C. L. Schneider, C. O. Gomes, R. P. King, J. A. Finch, D. Lin

Discussion is presented by Prof. King and Dr. Schneider of the paper by Lin, Gomez and Finch published in Trans.IMM C, vol.104, 1995, p.C155-C161. They conclude that the correction procedures used for simulated observations of simulated capped spheres are not reliable when applied to mineralogical data and that their relative ranking should not be extrapolated to other situations. Properly regularised solutions obtained with the use of appropriate kernels can be safely used for routine mineral liberation work and can be calibrated accurately against real mineralogical samples in the laboratory

Jun 19, 1905

By C. T. Aimone-Martin, V. L. Rosenhaim

The response of two structures to blast-induced ground vibrations were evaluated and compared in order to quantify the impact of different blasting operations, hard rock and soft rock blasting, upon the structures movements. The first instrumented structure was a residential type, located in the operational area of a coal mine, were blasting is conducted in soft rocks, such as siltstone and coal, with compressive strength ranging from 20 to 50 MPa. Blasting is characterized by low benches, with hole lengths averaging 5.5 m (18 ft) and low powder factors, 140 to 210 g/m3 (0.236 to 0.354 lb/yd3), resulting in very little horizontal movement of the rock during the blast. The second structure was a commercial building, located in a basalt quarry (compressive strength between 109 and 180 MPa), where blasting is conducted in 12 m (39.4 ft) benches with powder factors ranging from 200 to 350 g/m3 (0.337 to 0.590 lb/yd3). Velocity sensors were placed on the structure walls to measure wall displacements and calculate blast-induced strains in the walls, and these results were compared with the driving ground excitations. The calculated strains were compared with the limits of rupture of the cement grout used as mortar between bricks and as wall coating, considered to be the weakest material in the structure composition. The blast-induced response of existing cracks in the structures were also measured using displacement gages and compared to the daily influence of weather fluctuations, such as changes in temperature and humidity, on crack movement. The results show that for the same levels of ground motions the structures present different movement, resulting in distinct levels of induced strains on the walls depending on the geology. The influence of long-term temperature and humidity variations on crack movements on both structures is very similar, during the same seasons of the year. Crack response to long-term weather changes is 2.5 to 3.5 times greater than blast influences.

Jan 1, 2015

The Gympie Province is a unique tectono-stratigraphic unit which does not fit into the overall palaeogeographic pattern of the New England Fold Belt (NEFB), and part of it has been proposed as an exotic terrane more closely linked to New Zealand. From E to W the Gympie Province comprises 6 distinct units which can be evaluated as suspect or exotic terranes.

Jan 1, 1990

By Hiroaki Sakurai

Using batch reactors equipped with CO and CO2 gas sensor, CO oxidation over Au catalysts in room air has been investigated and compared with Pt catalysts. For both Au and Pt catalysts, the logarithm plot of CO concentration data during initial 10 min are well fitted by straight line, which mean the first order dependence on the CO concentration between 10 and 10,000ppm. The rate constant value (k) over Au and Pt catalyst generally decrease with increasing initial CO concentration above 100ppm. The value decrease especially rapidly between 100ppm and 1,000ppm over Pt catalysts, which should be due to the strong adsorption of CO on Pt. The k value over Au catalyst has become ten times or more as large as the Pt catalysts at 1,000ppm. Already for the initial concentration of 100ppm CO, the strongly adsorbed CO on Pt metal has been clearly observed by FT-IR during the reaction over Pt/Al2O3. A part of the strongly adsorbed CO has been remained on Pt after the disappearance of gas phase CO, and has gradually changed to gas phase CO2. Such strongly adsorbed species on Au has never been observed on Au/TiO2 under the same reaction conditions. These features of Au catalysts are favorable for indoor air purification.

Oct 1, 2003

By C. Liu, G. Zhu, Y. Liao

"The effect of particle size on the flotation kinetics of a low-rank coal was compared using air bubbles and oily bubbles. Five kinetic models were applied to the data from the tests to derive the relationships between the flotation rate constant, the ultimate recovery, and the particle size fraction. The results show that oily bubble flotation (OBF) resulted in a lower ash content and higher combustible recovery than air bubble flotation (ABF), especially for the -125 ??m size fraction. It was found that the flotation of a low-rank coal in individual size fractions using air bubbles and oily bubbles can be best described by a first-order model with rectangular distribution of tendency to float and the classical first-order model respectively. OBF yielded higher ultimate combustible and ultimate ash recoveries. The flotation rate constants for OBF were higher in all the models except for the fully mixed reactor model. The ultimate recovery values and modified rate constants first increased and then decreased as the size fraction decreased. The trend of the flotation rate constants as a function of the size fraction was similar in all the models except for the fully mixed reactor model. IntroductionFroth flotation is a physical-chemical separation method that is widely used in coal beneficiation for the separation of fine coal particles from associated minerals in water slurries. Since the cumulative recovery of a component in the concentrate is proportional to flotation time, the flotation process can be considered as a time-rate recovery process (Emad, Mohammad, and Bahram, 2010). Therefore, the general mathematical flotation models that incorporate both a recovery and a rate function can describe flotation time recovery profiles.The kinetics of coal flotation, as a measure of coal recovery as a function of time, have been studied in detail by many researchers. Different kinetic models have been proposed and published in the literature. The order and the rate constant are strongly dependent on the flotation conditions, such as the particle size and size distribution, the bubble concentration, the reagent concentrations, and the operating parameters (Zhang et al., 2013). The particle size has been reported to affect the flotation kinetics to a great extent (Yoon, Luttrell, and Asmatulu, 2002; Ntengwe and Witika, 2011). Numerous researchers have studied the aspects of flotation kinetics while paying special attention to particle size (Meftuni and Ibrahim, 1997; Rubinstein and Samygin, 1998; Graeme, 2012). However, the effect of particle size on the flotation kinetics of a low-rank coal has been relatively little discussed."

Jan 1, 2017

By H. Cevizci

"Numerical simulation of the plaster stemming method (PSM) was performed and compared with the conventional drill cuttings stemming method (DCSM). Many earlier in situ tests have proved that PSM can use the blast energy more efficiently than DCSM. Despite PSM generating more blast vibrations, it has advantages over DCSM such as better fragmentation and lower cost per unit volume of rock blasted. In this study, numerical simulation with Autodyn software using a 2D tool was employed to prove the efficiency of plaster stemming by comparing parameters such as pressure, Y-velocity, Y-force, internal energy, acceleration-Y, and compression. For example, the maximum pressure attained at the top of explosive column was 7 395 MPa for DCSM whereas it was as high as 11 945 MPa for PSM. Most of the computed parameters were significantly higher in PSM than those obtained for DCSM. This paper is the first study elucidating the efficiency of PSM by numerical simulation. It is concluded that PSM can save substantial amounts of money and effort. IntroductionThe stemming of blast-hole collars in surface mines with an inert material redirects blasting energy to the rock more efficiently, thus the blast-induced energy is utilized more effectively in breaking rock. With proper stemming, the gases should not escape due to loose stemming material. More efficient stemming with better confinement therefore increases the blast fragmentation. In addition, the distance of scatter is increased, giving rise to a looser rock pile that can be more easily loaded and transported. In order to improve the performance of a blast, stemming is used to help maintain the gas pressure over time (Konya and Konya, 2018). Proper stemming has been shown to increase the explosive efficiency by over 41% (Snelling and Hall, 1912).The most common stemming material in open pits and quarries is drill cuttings because of their ready availability at blast sites and the drill cuttings stemming method (DCSM) is a low-cost method. However, it has a major disadvantage in that dry drill cuttings eject very easily from blast-holes during an explosion, thus a great percentage of the blast energy is wasted and lost to the atmosphere. Cevizci (2012, 2014, 2013, 2017) studied blasting parameters in open pits and obtained better results with the plaster stemming method (PSM) in many limestone, basalt, and clay quarries. Moulding plaster is preferred for stemming because of the fast hardening time of 25–30 minutes."

May 1, 2019

By F. M. Lyon, F. J. Chernosky

Research efforts have not developed techniques for the complete desulfurization of coal that is needed to reduce air pollution caused by burning coal and to reduce the sulfur in metallurgical coke. Sulfur in coal occurs in three forms and of these, the pyritic sulfur occurs as discrete mineral particles and, at least theoretically, should be subject to physical rejection. Pyrite occurring in hard rock mining is concentrated by froth flotation; therefore, the possibility of removal of pyrite from coal by flotation does exist. With this possibility in mind, the flotation and adsorption characteristics of ore and coal-pyrite were compared while using potassium ethyl xanthate as the collector.

Jan 1, 1973

By W. U. Reimold, R. Armstrong, R. G. Charlesworth

For many years rockbursts have plagued the mining industry. Although much research has gone into studying this phenomenon only a small proportion of it has been on the actual source mechanism. A study was recently carried out to compliment the previous work done on the micro-structure of these features and to compare it to the micro-structures seen on shatter cones, from meteorite impacts. Both optical and electron-microscopy were used during the study. Samples were collected from rockburst sites by the CSIR Miningtek Rockburst Investigation Unit, and from the collar of the Vredefort Impact Crater. In the optical-microscopy study, the pinnate fractures observed in previous studies were seen in the shatter cones, through all the material. In the rockburst samples, the fractures were observed only in the quartz grains. Quartz micro-crystals were present in the rockburst samples and appeared to be the product of comminution rather than crystallisation. Cataclasite was present on rockburst and shatter cone fracture surfaces. The observations of the shatter cone surfaces with the scanning electron microscope are in agreement with previous work, including the observed presence of melt features, spherules and spheroidal material. The deformation in the rockburst samples is concentrated in the quartz grains, highly complicated with brittle features like sub-conchoidal structures and step structures present. A quartz spherule was identified on one rockburst sample and rhombic-dodecahedra type micro-crystals were present and seen fracturing off quartz grains. It is clear that the rockburst and shatter cone fracture surfaces are quite different; the shatter cone deformation is equally distributed across the sample and involves plastic-deformation, while the deformation in the rockburst material is concentrated in the quartz grains and, with the exception of spherule formation, is brittle. This study agreed with the current explanation for the formation of the rhombic-dodecahedra, and disproved the possibility of it being coesite or fine-grained meta-chert. Although there is no definite evidence of the mechanism of formation of the spherule, it is believed to be the product of vaporization and condensation, and the presence of this material indicates the extreme conditions under which this event occurred.

Jan 1, 2003

By Vincent Scovazzo

This paper compares the two most common pillar design equations used worldwide by John T. Boyd Company; the Mark-Bieniawski Equation, an empirical approach, and the Wilson Equation, an analytical approach. The equations are compared in two ways; by their stress profiles and by the Mark-Bieniawski Equation stability factor. Contrasting the two equations based on the visual comparison of their stress profiles is straight forward but evaluating their resulting pillar designs is problematic. Complications arise due to the differences between the use and development of stability factors for empirical equations and safety factors for analytical equations. To work around these differences, comparisons are made between the two equations by first calculating pillar sizes using the Wilson Equation and then determining the Mark-Bieniawski Equation stability factor for these pillars. Pillars designed using the Wilson Equation were based on site-specific data from 23 mines located in the Americas and Australia. Comparing sizes of pillars designed by the Wilson Equation to the Mark-Bieniawski Equation; the Wilson Equation pillar sizes were smaller in 77% of the cases, similar in 3%, and larger in 20% of the cases.

Jan 1, 2008

By Kim Le Roux

Cemented paste backfill (CPB) is a relatively new material that is increasingly being used in open stope mining across the world. The advantages offered by this high density tailings backfill include: a lower total cost associated with backfill placement, reduction of excess water underground; and the elimination of the need to process the tailings prior to placement. However, almost all backfill design is based on the performance of laboratory prepared samples since it is difficult to obtain undisturbed samples and conduct in situ testing due to the lack of access to backfilled stopes and the safety concerns in accessible areas. The logistical problems associated with conducting in situ testing and obtaining undisturbed samples of cemented paste backfill (CPB) has lead to the reliance in the mining industry on performance data obtained from laboratory prepared CPB samples. The degree to which these laboratory prepared samples represent the actual in situ condition is not well defined, partly because the influence of the mixing, transport and placement process on the material properties it is not well understood. Therefore, there is a need in the industry to understand how the material properties and hence the performance of the laboratory prepared CPB compares with in situ material. This paper reports the findings of an investigation into the in situ properties of CPB at a gold mine in Northern Ontario. A combination of in situ and laboratory testing, conducted on undisturbed samples, was used to quantify the in situ properties and performance of the CPB. These results were then compared to those of a laboratory based investigation of the same CPB. It is evident from this investigation that the bulk properties of the in situ CPB are different from those of the laboratory prepared CPB. The most significant observed differences are that the in situ CPB has a higher void ratio and a lower degree of saturation than the laboratory prepared material. The implication of these differences on the stability and performance of the CPB will also be explored in this paper. 1

May 1, 2002

By Anil K. Ray

This paper discusses in situ studies conducted at an Illinois Basin room-and-pillar coal mine to evaluate the performances of three different types of bolts under similar geologic conditions. The bolt types considered were the fully grouted passive rebar (FGPR), the resin-assisted mechanical anchor (RMAB), and the fully grouted tension rebar (FGTR). During the evaluations, the three types of bolts were installed in a single entry covering a span of three contiguous pillars with almost identical geologic conditions. The performances of the three types of bolts were investigated by instrumentting the bolts prior to installation as primary supports on cycle. Additional data on roof behavior was collected with extensometers and shear meters placed near the instrumented bolts. Excluding some time lapse right after installation, the performance evaluation of the three bolt systems was based on the data collected for over a year. Based on this study, it was found that the axial loads measured in the different bolt types did not show any identifiable patterns under the prevailing geological conditions at the mine. Therefore, the assumption that active bolts must be superior to passive ones is not always valid. Even though the studied mine did not have any thinly laminated, weak rocks in the immediate roof, the passive bolts performed as well as the active systems.

Jan 1, 2012

By Mark D. Hannington

Drilling of the active Trans-Atlantic Geotraverse (TAG) deposit during ODP Leg 158 indicated that the size of the mound-stockwork complex is approximately 3.9 million tonnes, including 2.7 million tonnes of massive and semi-massive sulfide (ca. 2% Cu) at the seafloor and 1.2 million tonnes of mineralized breccias (ca. 1% Cu) in a subseafloor stockwork. Quartz-pyrite veining in the stockwork zone extends from about 40 meters below seafloor (mbsf) to a depth of 95 mbsf. Siliceous wallrock breccias in the lower part of the stockwork grade abruptly into chloritized basalt breccias at the margins of the mineralized zone, and massive sulfides at the flanks of the deposit onlap relatively unaltered, partially hematized basalts. The pipe-like dimensions of the stockwork zone do not exceed the diameter of the sulfide mound. Comparisons with samples collected during earlier dive series confirm that the vent complexes at the surface of the mound are not representative of the bulk composition of the deposit. Dramatic vertical zonation of metals within the mound suggests that a long history of hydrothermal reworking has effectively stripped the constituents that are soluble at lower temperatures from the massive sulfides and concentrated them at the top of the deposit through a process of zone refining.

Jan 1, 1998

By T. M. Barczak

Stoppings are required to resist lateral forces on the face of the structure to control pressure differentials created by ventilating air. The design criteria in other parts of the world, including Australia, exceed this requirement by specifying over pressurization control to as much as 14 kPa (2 psi) in the active sections and 35 kPa (5 psi) in the main ventilating control areas. This is done to provide protection to the mine workers from explosive or air blast events within the mine that can create this over pressurization. The National Institute for Occupational Safety and Health (NIOSH) has been conducting research to develop a new testing protocol for rating mine ventilation stoppings. The premise of this work is that arch-loading conditions more accurately reflects the transverse load capabilities of stoppings in underground coal mines than the freestanding flexural strength test specified in the current Code of Federal Regulations (CFR). This work clearly shows that the transverse load capabilities of dry-stacked, mine ventilation stoppings are dependent upon the material strength of the block and the height and thickness of the wall, none of which are part of the current CFR criteria based on ASTM E-72 specifications. This paper compares the transverse load capabilities of several block materials and wall dimensions commonly used in stopping constructions based upon simulated three-hinge, rigid-arch loading tests of half-wall constructions in the NIOSH Mine Roof Simulator. The results indicate a wide-range of transverse load capacities, ranging from as low as 21 kPa (0.3 psi) to as much as 117 kPa (17 psi) for the conditions evaluated in this study. The current CFR criteria do not distinguish between applications of these stopping designs despite their wide variance in transverse load capabilities, which is not conducive to employing ventilation control strategies that isolate certain sections of the mine to prevent injury to mine workers.

Jan 1, 2009

By A. Y. Senchenko, Y. V. Kulikov

"The most common technique for selection of grinding equipment is the Bond method implying determination of a ball mill work index (BWi). The method is perfect for greenfield projects where a ball mill must be sized to grind an ore to 80% passing 75-100 µm. Finer grinding will require application of specific correction factors to a standard equation since initially the method was not designated for sizing of fine (minus 75 µm) grinding mills. Significant issues are related to selection of grinding equipment for treatment of originally fine materials when the standard procedure is not applicable. In 1984, Levin published his special technique titled as “A proposed test for the determination of the grindability of fine materials”, where he offered to use the standard Bond ball mill for determination of a specific grinding energy. However, this method did not meet with wide recognition. In 2017, Institute TOMS under guidance of JKMRC’s specialists performed a testwork aimed at verification of a new universal technique for determination of grindability of initially fine materials. An object was to test three aged tailings samples for the subsequent modelling of comminution circuit and evaluation of potential throughput of the existing ball mills with final grind size of ?80 passing 53 µm. In parallel with development of the universal approach, TOMS’ specialists carried out tests according to the Levin method in order to compare the data resulted from two methods. The results from lab ball mill testing on three products following the alternative test methods are presented in this paper along with the testing methodology and results analysis. Specific energy was calculated for regrinding of the materials to the required size. Difference of the examined methods and assessment of their applicability for sizing of full-scale equipment are explored."

Jan 1, 2018

By W. G. Pariseau

"BACKGROUND Strata mechanics of three underground coal mines in central Utah were recently studied using a specialized finite element technique [1, 2, 3, 4]. The purpose of these studies was to examine relationships between mining and related seismicity. The longwall mining method was predominant, but room and pillar mining and barrier pillar mining were also used in the mines studied. In every case, high correlation between face advance and seismicity was observed. The specialized finite element technique allows for whole mine subsidence estimation at a kilometer scale and for stress concentration at a meter scale near a working face. Comparisons of observed mine subsidence with model subsidence allows for model verification and justifies detailed analysis of the evolution of stress at seam level. Yielding is expected near a longwall panel face and as the face is advanced the extent of yielding is expected to increase. The total number of element failures in a finite element mine model also increases, so a high correlation with associated seismicity, say, event count, is not so surprising. A brief review of recent coal mine studies in Utah is presented in the sequel. The procedures are given in some detail in application to trona mining. COAL MINES The analysis procedure begins with the geologic column and proceeds to specification of strata properties including elastic moduli and strengths. Joints are an integral part of the geologic model and joint set specifications include dip direction, dip, spacing, stiffness and strength. Spatial variability of strata properties in situ may be taken into account during analysis. Topography is important, especially to Utah coal mines that are developed from outcrop in regions of high relief. Mesh generation follows specification of data input. Finite element analysis is next, followed by reporting of results. In this regard, all finite element computations are based on first principles. No model “calibration” is done that involves adjusting computer input to obtain a fit to mine measurements such as subsidence profiles. Figure 1 illustrates the concept of a dual node – dual mesh finite element concept. Dual nodes are roof and floor nodes in mined areas. These nodes replace the seam. The dual mesh is a much refined mesh with displacement boundary conditions obtained from the big mesh and thus replaces the effect of mining in the region outside the dual mesh. The refined dual mesh allows for details of stress, strain and displacement near a working face of interest. Strata Properties Strata properties include elastic moduli and strengths of intact rock between joints and joint stiffness and strength for each joint set present. Both properties sets may be obtained from laboratory test data, although testing joints poses a sampling problem in the field. The statistical variability observed in laboratory test data, say, in unconfined compressive strength tests is often characterized by much variation. In fact, the coefficient of variation, ratio of standard deviation to mean, is often greater than 35 percent. When sample values are mapped back to the mine location, one may observe spatial variability as properties vary from place to place as shown in Figure 2. Spatial variability may be incorporated into a finite element analysis with probabilistic assignment of properties to elements in accordance with measured variability. The result is a more realistic model of rock in situ."

Jan 1, 2017

By Deon van Tonder, Chuck Edwards

"Solvent extraction is used in many acid leach uranium circuits to upgrade and purify the uranium bearing solution prior to final product precipitation. A number of strip options are available to recover the uranium from the loaded organic. This paper compares some of these options on the basis of typical capital and operating costs for both the solvent extraction circuit and downstream uranium processing, and also considers other related issues. The aim is to provide an objective comparative assessment of these options that could be used as a qualitative guide in process selection or improvement.INTRODUCTIONIn the 1950s the uranium industry was the first industry to adopt solvent extraction (SX). Since then, the flow sheets have advanced. Notable was the introduction of tertiary amine extractants by General Mills during the late 1950s and 1960s (Mackenzie, 1997; Mackenzie, 2005), which is still today considered ‘state of the art’ technology. Initially solvent extraction was done in combination with ion-exchange, such as in the Eluex flow sheet, followed later by the adoption of direct solvent extraction from pregnant leach solution (PLS). From a strip-precipitation perspective ammonia strip followed by ammonium diuranate precipitation was widely adopted as the norm in the uranium industry. More recently alternative strip systems have been receiving increasing attention due to the environmental, health and safety issues associated with ammonia and the resulting calcined uranium oxide product. This paper explores some of these alternative options in comparison with the ammonia route.It is important to acknowledge that each operation is unique, and therefore the information and philosophy outlined in this paper serve to provide a general guide that can be potentially applied and adapted to any site or project."

Jan 1, 2013

By D. Van Tonder, C. Edwards

ABSTRACT Solvent extraction is used in many acid leach uranium circuits to upgrade and purify the uranium-bearing solution prior to final product precipitation. A number of strip options are available to recover the uranium from the loaded organic. This paper compares seven of these options on the basis of typical capital and operating costs for the solvent extraction circuit and downstream uranium processing and also considers other related issues. The aim is to provide an objective comparative assessment of these options that could be used as a qualitative guide in process selection or improvement.

Jan 1, 2013

By B. Boberg

Annual requirements of uranium to fuel the 441 currently operating nuclear reactors in the world is 170 million pounds U3O8 (65,385 tonnes U). 2005 global production was 108 million pounds (41,595 t U), of which only 2.7 million pounds U3O8 (1,039 t U), was produced from US uranium deposits. US nuclear utilities represent 26% of the global nuclear capacity but only obtain 2.5% of their supply from US uranium mines. Since 2002 annual global production increased from 93 million pounds U3O8, a 15% increase and the US production increased from 2.3 million pounds to 2.7 million pounds, a 17% increase. Reasonably Assured Resources (RAR) of 12,332 million pounds U3O8 (4.7 million t U) are reported from a wide variety of countries and types of uranium deposits. This RAR suggests that there are significant global uranium resources potentially producible at a price of $50 per pound U3O8 ($130 per kg U). Table 1 shows the distribution of global uranium resources. This table, prepared from historic data maintained by the World Nuclear Association, does not include some areas of the world that are currently considered to contain significant resources of uranium such as Mongolia. Table 1 suggests that the United States contains the fourth largest uranium resource in the world with 890 million pounds U3O8, behind Australia, Kazakhstan and Canada in resources that are producible at a price of $50 per pound U3O8 ($130 per kg U). Within the US, the states that demonstrate the greatest amount of uranium resource are Wyoming, New Mexico, Texas and the Four Corners Area of Arizona, Colorado and Utah.

Jan 1, 2007

By A B. Christie, R L. Brathwaite, A Swain

This study provides quantitative data on vein distributions within seven selected epithermal systems, four deposits hosted predominantly in rhyolitic rocks: Ohui, Wharekirauponga, Neavesville and Komata, and three major deposits mainly hosted in andesite: Martha Hill, Golden Cross and Karangahake. The technique of fractal analysis was used to help quantify vein distribution patterns. The fractal dimensions (D) obtained are related to the number of veins per metre, the degree of vein clustering, and vein thickness. An increase in any of these tends to increase the fractal dimension.   It is the relative relationship of these parameters that is important for the characterisation of vein distributions, and the comparison of different deposits. Within the same systems, the rhyolites are characterised by lower vein thickness and vein density, but have a higher degree of vein clustering compared to the andesites. Sections that are deeper and more centrally located within the epithermal system and/or close to major structures tend to have a greater degree of vein clustering, higher average vein thickness and density, and higher thickness to density ratios. This depth control is most pronounced in epithermal gold-silver systems because they have formed at shallow crustal levels where physico-chemical conditions change drastically both vertically and laterally. In addition to lithological and structural control at vein scale, deposit scale structure is of great importance in the deposition of economic mineralisation. Large open vein structures capable of sustaining high fluxes of hydrothermal fluid, favourable for the formation of large, high-grade gold deposits, are best developed in andesite sequences.

Jan 1, 1995