Search Documents

By H. E. Bartlett

Previously, the grade of ore at the shaft head was taken to be equivalent to the grade measured in the faces by means of chip sampling. The tonnage mined from the stopes is determined from survey volumetric measurements. Estimates of shaft head grade and tons include the grades and tons of waste, development ore, sweepings from old areas, and other sources, not all of which are sampled. The tonnage of ore delivered at the shaft is often estimated using skip factors. These factors are subject to manipulation and not all the skips are filled to the same level. Consequently, both the grade and tonnage of ore delivered at shaft head are subject to a measure of uncertainty. For many years, these uncertainties have led to disputes between shaft managers and metallurgists over the reconciliation between shaft tons and grade and plant tons and grade. These problems are compounded when several shafts feed a central processing plant. Shaft bonuses and a shaft?s profitability are affected by poor gold allocation methods. Without measurement of the grade and tonnage from all shafts, a poorly performing shaft could benefit at the expense of the other shafts. In order to improve the measurement of grades and tonnage at the shaft head, which is the custody transfer point between the mine and surface operations, South African mining companies have developed a methodology to measure the mass of ore from each shaft using rail weighing equipment, and the run-of-mine grades delivered from each shaft using cross-stream (go-belt) sampling. Harmony Gold Mining Company has implemented the rail weighing and go-belt sampling methods for their shafts. This paper deals with the statistical tests on the procedures and control measures that are in place to establish confidence in the gold allocations. The metal accounting system for examining the allocations follows the first principle of the AMIRA P754 Metal Accounting Code of Practice namely ?The metal accounting system must be based on accurate measurements of mass and metal content. It must be based on a full check-in check-out system?. The check-in is mass and grade of run-of-mine ore and the check-outs are the plant input grades and the recovery and residue from the plant. The balances between check-in and check-out are the subject of the statistical analyses, which have been conducted using the classical statistical tests and procedures that are generally used in all fields of endeavour. It is concluded that the gold allocation procedures that rely on go-belt sampling and rail weighing of ore from shafts has been successfully implemented, and that the system, which has been in operation at Harmony since 1993, has won general acceptance from stakeholders, with the gold allocations being perceived as being fair to all. Apart from the on-mine issues of bonuses and fair allocations, accurate metal accounting assists in the broader issues of generating accurate production figures for management. An added benefit of weighing and sampling ore at the shaft head has been that shafts now have daily moving average grades, which are used for grade control. The go-belt sample at the shaft head is the first accurate sample of production in the value chain from face sampling to gold bullion. Keywords metal accounting, go-belt sampling, run-of-mine sampling, codes of practice, AMIRA P754, train weighing, statistical analysis, gold split, allocation of gold, metal balance, grade control.

Jan 1, 2014

By R. H. Seraphim

"Western Mines Limited produces gold, silver, copper, lead and zinc from a number of Paleozoic volcanogenic deposits near Buttle Lake on Vancouver Island. The deposits were emplaced along a series of submarine linear volcanic ridges that extended for at least 6,000 metres in a northwesterly direction.Rock types mapped include: andesitic, dacitic and rhyolitic flow s and volcanic breccias; detrital breccias, tuffs and hematitic silts; black and green cherts; and a variety of dykes and sills, including quartz porphyry and diorite. Orebodies are associated invariably, and in places intimately, with quartzsericite schist and rhyolitic rocks. The ore is complex, varying in mineralogy and tenor both within each deposit and from one deposit to another.Recent studies have increased our knowledge of the structure and left no doubt that the orebodies were deposited both within the throats of the linear vents and on paleoslopes adjacent to these vents.IntroductionWestern Mines Limited operates an 850-tonne-per-day mine, located 85 km south of Campbell River, B.C. The mine is at approximately 400 metres elevation, near the south end of Buttle Lake on Vancouver Island (Fig . I). It has operated since early 1967 and produced 3,650,000 tonnes with an average grade of 1.9 g (0.06 oz) gold, 93 g (3.0 oz) silver, 1.6 per cent copper, 1.0 per cent lead and 7.5 per cent zinc per tonne by the end of 1978. Current reserves are 1,157,000 tonnes of 2.7 g (0.09 oz) gold, 117 g (3 .8 oz) silver , 1.2 per cent copper, 1.2 per cent lead and 8.0 percent zinc per tonne (December 31, 1978).Ore from one zone, known as Myra High-Grade, is mined and milled separately because of its grade. Ore milled prior to December 31,1978, plus reserves, 180,715 tonnes of6.6 g (0.21 oz) gold, 554.5 g (17 .7 oz) silver, 0 .8 per cent copper, 3.2 per cent lead and 10.6 per cent zinc are included in the figures listed in the preceding paragraphs."

Jan 1, 1980

By R. W. Amstutz

Abstract. The shaft was drilled on U.S. Government land in the Piceance Creek Basin and was funded by the U.S. Bureau of Mines. It was drilled to 723 meters (2 371') to provide access to rich deposits of oil shale, nahcolite and dawsonite for geotechnical test work and environmental mining research. Rotary drilling and dual string, reverse-air, jet-assist circulation methods were used. Drilling required 8 months. Site preparation, running and cementing the casing required another 4 months. A 2.44-meter (96-inch) ID steel welded casing, 5.1 cm (2 inches) thick was installed and 2 700 tonnes (63 000 sacks) of cement were slurried and pumped into the annulus.

Jan 1, 1980

By A Gerson, S Miller, R Smart, J Thomas

Environmental Geochemistry International (EGi) and the Ian Wark Research Institute (IWRI) collaborated in a three-year Australian Minerals Industry Research Association (AMIRA) research project (Project Number P387A) focussed on the prediction and kinetic control of acid mine drainage. As part of this project, single addition, and sequential net acid generation (NAG) testing was carried out on sulfide mineral samples. The test samples comprised a range of synthetic waste materials prepared using a quartz bulk and controlled quantities of the commonly occurring sulfide minerals; pyrite (FeS2), pyrrhotite (Fe7S8), arsenopyrite (AsFeS), chalcopyrite (CuFeS2), chalcocite (Cu2S), covellite (CuS), bornite (Cu5FeS4), galena (PbS) and sphalerite (ZnS). NAG test results clearly show that there are significant differences between the acid forming potential of sulfide minerals. The NAG test results for each mineral were compared with theoretical acidity values (NAPP*). The NAPP* calculation was based on stoichiometric oxidation reactions assumed for each of the sulfide minerals. Sequential NAG values for pyrite corresponds closely to predicted NAPP* values, confirming the assumption that pyrite oxidation occurs according to the stoichiometry assumed. For the pyrite sample with 5 wt per cent S, two NAG stages were sufficient for all of the sulfur to be released to solution. Since sulfides in waste materials will generally be dominated by pyrite, results also confirm the validity of including NAG testing for ARD evaluation of most waste rock. NAG testing confirms that of the sulfide minerals examined, only pyrite, pyrrhotite, arsenopyrite and chalcopyrite are acid producing, ie able to produce leachate with a pH less than 4.5. In the case of sphalerite, galena, bornite, chalcocite and covellite the pH of the final NAG solution was greater than 4.5. The significant differences between the acid forming potential of different sulfide minerals could lead to overestimation of acid producing potentials based on NAPP values calculated from total per cent S alone, ie assuming all S is present as pyrite. The oxidation of galena during the NAG test resulted in the consumption of acid. The examination of the galena surface by X-ray photoelectron spectroscopy (XPS) after oxidation in peroxide solution indicated the presence of polysulfide. The reaction of sulfide to polysulfide is acid consuming, thus explaining the net acid consuming behaviour of galena in the NAG test. This reaction pathway may provide a possible explanation for the lower than expected acid production observed for some of the other sulfide minerals examined.

Jan 1, 2003

By Lloyd M. English

Accurate knowledge of the gas content of coal seams has always been important, according to the work of W. P. Diamond and J. R. Levine, 1981, but is becoming increasingly important for several reasons: Excessive gas emissions in the mine can create explosion hazards. Depletion of shallower coal reserves is forcing underground operations to greater depths where coal gas content is often greater. Natural gas prices and the need for new energy sources make naturally occurring coal gas a potential resource of economic interest. Information on coal gas content will impact mine designs, mine ventilation system specifications, and the effectiveness of methane control procedures used. Accurate resource assessment is needed for projecting and designing potential gas utilization enterprises. The emphasis upon methane hazards in mining operations, and upon potential methane utilization, has led us to improve upon the procedures used to determine coalbed methane levels.

Jan 1, 1984

By Yingbao Yang, Jinyue Liu, Yonggang Wei, Bo Li

Due to the depletion of copper ore with high-grade and low impurity contents, the raw materials for copper smelting exhibit low copper contents and high impurities. High-impurity elements lead to the poor performance of the final copper product; thus, the removal of impurity elements during the oxygen-enriched copper smelting process is crucial. In this study, based on the production of high-grade matte, the occurrence states of impurity elements (such as As, Pb, and Zn) are analyzed in the input and output materials of the oxygen-enriched smelting process, and the phase transformation mechanisms of impurity elements are revealed during the oxygen-enriched smelting process. The effects of temperature, Fe/SiO2 ratio and oxygen concentration on the distribution behaviors of impurity elements in dust, slag and matte and the removal rates of impurity elements in matte are clarified. The results show that the impurity elements mainly exist as As4, As4S4, As2O3, Pb, PbS, PbO, Zn and ZnS in dust, as FeAsO4, As2O3, FeAsS, PbSiO3, Zn2SiO4 and ZnFe2O4 in slag, and as As, FeAsS, PbS, PbO and ZnS in matte. High smelting temperature, low Fe/SiO2 ratio and low oxygen concentration are beneficial for improving the matte grade and the removal rates of impurity elements Pb and Zn in matte. Low smelting temperature, low Fe/SiO2 ratio and high oxygen concentration are beneficial for removing impurity element As in matte.

Aug 8, 2023

By Walter N. Campbell

The Simon Silver Mine is located near Mina, Nevada, approximately 50 miles northwest of Tonopah and 40 miles east of Hawthorne. It is an old property which was last worked in 1932 and closed operations due to problems arising from low metal prices. Federal Resources Corp. acquired the property in 1963 and undertook a completely new development program since the old surface plant and underground workings were beyond feasible repair. The deposit of silver, lead, zinc and copper is predominantly replacement of limestone. In the summer of 1964, a new surface plant was constructed and on December 2, 1964, sinking of a three compartment vertical shaft started and was completed to a depth of 1088 feet in April of 1965. Following the sinking program an 850 foot crosscut was driven to intersect the mineral zone on the 1000 level after which drifting, crosscutting and raising totaling some additional 4280 feet has been completed along with considerable diamond drilling and long holing into the walls. To date, two ore bodies have been encountered and test stoping operations are underway together with development of the 800 level for servicing, ventilation and access. The present development of ore is directed towards substantiating sufficient reserves to warrant the construction of a flotation mill.

Jan 1, 1967

By Rob D. Morrision, Paul W. Cleary

"The best accepted method for design of autogenous and semi-autogenous mills is to carry out pilot scale test work using a 1.8 m diameter by 0.6 m long pilot scale test mill. The load in such a mill typically contains 250,000-450,000 particles larger than 6 mm, allowing correct representation of more than 90% of the charge. This scale of DEM model is now within the practical range of standard workstations running an efficient DEM code.This paper describes various ways of extracting collision data from the DEM model and translating it into breakage and wear estimates. Account is taken of the different breakage mechanisms (impact and abrasion) and of the specific impact histories of the particles in order to assess the breakage rates for various size fractions in the mills. These impact histories have the potential to be directly related to breakage as they most resemble the single particle breakage tests which are used by the JKMRC to calibrate more traditional population balance models for variations in ore behaviour. However, the specific impact histories also show a background of many low energy collision events within the mill. This Background may constitute what is usually regarded as “abrasion” breakage. As might be expected, high impact events are concentrated in the toe of the charge and near to the mill shell where the forces associated with the load bearing on each particle is more concentrated."

Jan 1, 2003

By P J Savory

Ranger is a uranium mining operation located within one of the most environmentally sensitive regions of Australia. Geology has played an important role in mine planning, operations and rehabilitation since discovery of the deposits 25 years ago. The Ranger orebodies have a reserve total of 166 300 tonnes of U308 including 90 400 tonnes of U308 at North Ranger. At current milling rates these reserves give a mine life greater than 40 years. The uranium is hosted in Lower Proterozoic rocks of the Cahill I ormation and is associated with strong chlorite alteration. In pit grade control involves the use of calibrated probes and a radiometric truck discriminator allowing fast, accurate grade determination of material. Ore is defined at a cut-off grade of 0.2 per cent U308. Mineralised material is defined by the Uranium Mining (Environmental Control) Act 1979 (UMEC Act) at a grade of 0.02 per cent U30s. Natural water levels in the #1 Pit area vary as much as 7.5 metres between wet and dry seasons. Ground water inflow into the pit during the dry season is 600 m3/day. No significant slope failures have occurred in the pit. The environmental impact of mining operations is constantly monitored. Rehabilitation of non-mineralised stockpiles is ongoing. This paper discusses the interaction between geology, hydrogeology, rock mechanics and grade control with the present mining operation.

Jan 1, 1994

By Milance Mitovski, Milorad Cirkovic, Vlastimir Trujic, Dragana Živkovic

"In this article, it is analysed the possibilitis and the effects utility waste energy of process sulfuric acid production in the existing technical-technological conditions. The utility second process energy, reduction the fuels consumption, reduction environment pollution and its effects are shown, too. High heat quantity is produced as the result of exothermal reactions in the production process of sulfuric acid in Bor. A part of this heat is sufficient for the process development, and significant part of this heat presents low-temperature heat energy that is irretrievable lost. Investigations were carried out based on production of 485 t of monohydrate/day of sulfuric acid. By recovery of this heat quantity for heating purposes, about 20.000 – 25.000 t of coal/year would be saved, what would significantly improve the power efficiency of the smelter copper production. IntroductionTechnological process of copper production in RTB BOR (Serbia) produces a significant quantities of high temperature secondary (waste) heat energy that is uncontrolled “lost” in the vicinity. Secondary heat energy is produced in pyrometallurgical copper production in quantity of 1,621 kWth/kg of dry concentrate or existing heat (that presents a difference of heat at furnace outlet and heat at temperature over temperature of dew point of sulphuric acid gases) in quantity of 1,315 kWth/kg of dry concentrate. For designed capacity of Copper Smelter in Bor, the existing heat power of secondary energy is 92,111 MWt, what is an equivalent to the power of steam-turbo plant where the Rankine-Clausius heating cycle is done, and that is 28,334 MWe."

Jan 1, 2008

By John Griffen

This paper records over twenty years' experience with the use of the Rheolaveur system in the United States, showing its ability to meet changing conditions caused by the dirtier mine output of present-day mechanical mining methods. Data are given on size ranges handled, number and capacity of units installed, maintenance experience, and operating results on two-product and three-product separations. IT is not the purpose of this paper to discuss the principles employed in the Rheolaveur system of fine-coal cleaning as these have been fully covered in the technical literature of the Institute."2 Rather, our purpose will be to record the highlights of over 20 years' experience in the United States with Rheolaveur fine-coal launders, which will indicate their capabilities and costs of cleaning and their adaptability to meet the changing conditions caused by present-day mechanical mining methods. Rheolaveur fine-coal launders are used in the United States to treat a wide variety of size ranges of fine coal, the coarsest being 1/2 in. to 0 and the finest about 1/8 in. to 0. In Europe feeds as coarse as 5/8 in. and as small as 1/2 mm (28-mesh) to 0 are being cleaned. The size ranges usually handled in this country are 3/8 to 1/4-in. round to 0. Thirty five units with a combined hourly feed capacity of 3200 tph have been installed in the United States. Individual units are cleaning as little as 25 tph, while others are cleaning as much as 200 tph. Rheolaveur fine-coal launders offer cleaning units of high capacity and are outstanding in requiring a minimum of building space per ton of input. Experience has shown that maintenance costs are low. Several installations were made in the Pittsburgh district from 1928 to 1930, and since that time operation has been largely two shifts per day. Minor repairs to liner plates and Rheo boxes have been required during the intervening years, but no major replacement of launders was required until 1947 and 1948. At one of these plants, launders operated for 17 years before they were replaced and during that period over 26,000,000 tons of —5/16-in. coal were cleaned. The effect of the much dirtier raw coal produced by mechanical loading of the Pittsburgh seam is shown by the following tests. The data in table I summarize the performance of a Rheo fine-coal unit when cleaning hand-loaded raw coal. The —4-in. raw coal fed to the cleaning plant analyzed only 8.16 pct ash and 1.31 pct sulphur and contained 4.6 pct sink, 1.55 sp gr. Two years later, in 1945, mechanically-loaded coal from the same mine was cleaned in the same plant. The feed, —4 in., then analyzed 20.10 pct ash and 1.53 pct sulphur and contained 20.2 pct sink, 1.55 sp gr. The performance of the Rheo fine-coal unit when cleaning this coal is given in table 11. It will be noted that the +48-mesh coal is cleaned almost as thoroughly as in table I. The large amount of high-ash slimes produced from the dirtier feed are responsible for the higher ash of —--48-mesh washed coal in table 11. In this plant the Rheo fine-coal unit did not get the full load of refuse resulting from mechanical

Jan 1, 1951

By A. F. Agnew

Discussion by A.E Agnew Strachan's paper summarizes a wealth of data on water-storage dam "incidents" published by USCOLD and ASCE in 1988, together with a review of tailings-dam incidents published by USCOLD in 1994. The data is summarized by text and 11 histograms. The histograms present a useful compilation of these data, although the patterns within the histogram blocks did not survive the Mining Engineering printing process - so their usefulness is impaired. Unfortunately, the text discussion is too cursory and does not highlight the significant relationships among the various causes of the incidents. Four paragraphs of text and Fig. 1 were devoted to water-storage dams, but the discussion is too sketchy, and Fig. 1 is unclear. Important statistics that could be uncovered are: •the total number of dam incidents tabulated were 500+, •earth embankments constituted 73% of those in operation and •incidents related to earth embankments were 75%, of which dam failures accounted for 24%, major repairs accounted for 42% and accidents accounted for 34%. Strachan concluded that "the majority of the dam failures are associated with overtopping (p. 49), and that the majority of incidents are associated with spillway and other facility structures" - although this is difficult to discern from Fig. 1. Confusion regarding Strachan's use of the terms "dam failures" and "incidents" is compounded by his use of "failures" and "accidents" in one paragraph of his Introduction section followed by his use of "incidents" in the next paragraph. Getting to the meat of Strachan's important paper, the tailings-dam incident survey (USCOLD, 1994), the connection between Figs 2 through 11 seems obscure. First, in the four paragraphs of introductory material, one learns that "incidents consisted of failures, accidents, seepage and other incidents of performance that did not meet design criteria" (p. 50). One does learn the helpful caveat that "structures not related to mill tailings were not included in the review" - so that the Buffalo Creek (West Virginia in 1972) and Aberfan (Wales in 1966) coal-mine failures, together with the ongoing metal-mine remediation work at Summitville (Colorado) do not muddy the waters of his paper. The all-too-brief discussion of the data shown in Figs. 3 through 11 (185 dams in USCOLD plus 26 dams in UNEP) reveals the following: •Figure 3: 57% of the incidents (failures and accidents) occurred with dams less than 20-m (66-ft) in height. •Figure 4: 63% of the incidents were due to failures alone (subset of Fig. 3). •Figure 5: Active dams had 82% of the accidents and 93% of the failures. •Figure 6: Types of dams (upstream, water-retention, downstream, centerline and "unknown") separated according to incidents. •Figure 7: Breaks out the incidents in Fig. 6 by failures, accidents and ground water. •Figure 8: Breaks out incidents by type (overtopping, slope stability, earthquake, foundation, seepage, structural, erosion, mine subsidence and "unknown"). •Figure 9: Separates Fig. 8 data into active and inactive dams. •Figure 10: Separates Fig. 9 data into failures and accidents. •Figure 11: Same as Fig. 8, but broken out by tailings-dam type (Fig. 6). Figures 6 and 7 show that most of the failures are associated with upstream or "unknown" types. Figure 9 shows that the leading causes of incidents for active dams

Jan 1, 2003

By John Griffen

This paper records over twenty years' experience with the use of the Rheolaveur system in the United States, showing its ability to meet changing conditions caused by the dirtier mine output of present-day mechanical mining methods. Data are given on size ranges handled, number and capacity of units installed, maintenance experience, and operating results on two-product and three-product separations. IT is not the purpose of this paper to discuss the principles employed in the Rheolaveur system of fine-coal cleaning as these have been fully covered in the technical literature of the Institute."2 Rather, our purpose will be to record the highlights of over 20 years' experience in the United States with Rheolaveur fine-coal launders, which will indicate their capabilities and costs of cleaning and their adaptability to meet the changing conditions caused by present-day mechanical mining methods. Rheolaveur fine-coal launders are used in the United States to treat a wide variety of size ranges of fine coal, the coarsest being 1/2 in. to 0 and the finest about 1/8 in. to 0. In Europe feeds as coarse as 5/8 in. and as small as 1/2 mm (28-mesh) to 0 are being cleaned. The size ranges usually handled in this country are 3/8 to 1/4-in. round to 0. Thirty five units with a combined hourly feed capacity of 3200 tph have been installed in the United States. Individual units are cleaning as little as 25 tph, while others are cleaning as much as 200 tph. Rheolaveur fine-coal launders offer cleaning units of high capacity and are outstanding in requiring a minimum of building space per ton of input. Experience has shown that maintenance costs are low. Several installations were made in the Pittsburgh district from 1928 to 1930, and since that time operation has been largely two shifts per day. Minor repairs to liner plates and Rheo boxes have been required during the intervening years, but no major replacement of launders was required until 1947 and 1948. At one of these plants, launders operated for 17 years before they were replaced and during that period over 26,000,000 tons of —5/16-in. coal were cleaned. The effect of the much dirtier raw coal produced by mechanical loading of the Pittsburgh seam is shown by the following tests. The data in table I summarize the performance of a Rheo fine-coal unit when cleaning hand-loaded raw coal. The —4-in. raw coal fed to the cleaning plant analyzed only 8.16 pct ash and 1.31 pct sulphur and contained 4.6 pct sink, 1.55 sp gr. Two years later, in 1945, mechanically-loaded coal from the same mine was cleaned in the same plant. The feed, —4 in., then analyzed 20.10 pct ash and 1.53 pct sulphur and contained 20.2 pct sink, 1.55 sp gr. The performance of the Rheo fine-coal unit when cleaning this coal is given in table 11. It will be noted that the +48-mesh coal is cleaned almost as thoroughly as in table I. The large amount of high-ash slimes produced from the dirtier feed are responsible for the higher ash of —--48-mesh washed coal in table 11. In this plant the Rheo fine-coal unit did not get the full load of refuse resulting from mechanical

Jan 1, 1951

By D. Jenkins

THE Franklin Laboratory was designed mainly for the analysis of the products from the two concentrating mills situated at Franklin and Sterling Hill, the most important determinations being the zinc, iron, manganese, lime, moisture and silica contents of the ores. Within the last 4 years, however, it has been found advisable to examine most of our supplies, and therefore additional space and equipment was allotted for the examination of such substances as oils, greases, soaps, alloys, fuels, paints, explosives and water. The building as shown in elevation (Fig. 1) and plan (Fig. 2) consists of a main or routine laboratory, an experimental laboratory, a water FIG. 1.-WEST ELEVATION. analysis room and the other accessories such as stock room, library, sampling and drying rooms and office. The main laboratory is further divided into a balance room and zinc analysis room. As the greater part of our determinations are volumetric, good light is a very important factor and with this in view the interior walls of the building were constructed of white enameled brick laid in Keene cement. These also have the additional advantage that they give a clean and attractive appearance to the room. The windows are large, being 76 by 44 in. and even on very dark days the light is good throughout the whole laboratory. Our zinc titration requires a constant light and as, on many days, especially in winter, the light is variable, we installed two Artificial

Jan 8, 1917

By W. S. Weldon, R. M. G. Machunter

The use of spiral separators for the beneficiation of fine coal typically in the size range - 2+0.lmm is now well established process technology. Developments during the mid 1980's resulted in the production of cost effective, efficient large diameter coal spiral separators. Recent developments and research have been geared to optimise metallurgical performance, materials handling and equipment geometry of large diameter coal spiral separators.

Feb 24, 1997

By Jaques A. L, Heinrich C. A

Mt Todd lies towards the south end of the Pine Creek Geosyncline, sandwiched between the Cullen Batholith and the Edith Falls Basin, in an area of regionally metamorphosed sandstones, siltstones and shales. The Batman deposit, some 3 km west of Mt Todd is a sheeted quartz vein system hosted within greywackes of the Palaeoproterozoic Burrell Creek Formation (Finniss River Group). The gold-bearing quartz veins, trending 010¦-020¦ and dipping 60¦-70¦E, are commonly 0.5 to 5 cm wide and occur at intervals which are from 10 cm to 1 m apart. Gold tends to be restricted to the quartz veinlets and immediately adjacent country rock. Grade Control utilises two 4 m 'collar' samples which are collected from each 8 m production blasthole. The irregular distribution of the quartz veins and erratic gold mineralisation has resulted in a disseminated 'salt and pepper' type of gold occurrence, with only a poorly defined footwall. Delineation of ore and waste blocks for mining uses an approach which combines the smoothing effect of kriging with a pragmatic view of the raw data. Large ore blocks with simple geometry help to minimise mining dilution, reduce block variance and give good reconciliations to interpreted ore tonnes and grade. Daily mine face advance is reported against truck counts and provides continual comparison with budgeted production. All data processing uses Gemcom software, together with Lotus spreadsheets.

Jan 1, 1994

By C A Heinrich, Jaques A. L, L A I Wyborn

Most orebodies have cross-sections of less than I km2 and hence do not offer a particularly large target for exploration. Fortunately, although in the geological record ore deposits are small and rare and result from the exceptional coincidence of certain geological processes, these processes are mappable on a much larger, district to regional scale and constitute a mineral system in which the ore deposit is the central feature. A mineral system can therefore be defmed as 'all geological factors that control the generation and preservation of mineral deposits, and stress the processes that are involved in mobilising ore components from a source, transporting and accumulating them in more concentrated form and then preserving them throughout the subsequent geological history'. The mineral system concept emphasises that for many ore deposit types, although economically viable mineralisation may only occur on a scale of say, hundreds of metres, the total system offluid-rock interactions that led to ore formation can extend over a distance of tens to hundreds of kilometres around the deposit. When mapped out, the total mineral system provides a far larger exploration target than the actual ore deposit itself. Important geological factors defining the characteristics of any mineralising system include: I. sources ofthe mineralising fluids and transporting ligands; 2. sources ofthe metals and other ore components; 3. migration pathway; 4. thermal gradient; 5. energy source; 6. a mechanical and structural focusing mechanism at the trap site; 7. chemical and/or physical traps for ore precipitation. Many of these factors individually are common through

Jan 1, 1994

By J. H. East

"INTRODUCTION Diesel locomotives have been used successfully in underground mines of Europe for 18 years, and much has been written on the technical details of the design of these machines and the safeguards provided by the manufacturers. The actual conditions under which' the Diesel locomotives are used have not been publicized as well and consequently are not well-known. In 1945 the writer observed numerous Diesel locomotives working underground in coal and metal mines of several European countries and was favorably impressed with the practical approach to the safe use of this- type of equipment.Recent statistics showing the number of Diesel locomotives in use are not available now because of the World War, but the extent to which this type of mine locomotive was used immediately before the war is indicated by table 1, which classifies the number of locomotives used in several countries according to the type of power, such as compressed air, Diesel, and electric.Mining conditions observed by the writer in the European countries are very different from those in most coal-mining districts of the United States. The surface plants are larger and are permanently constructed of brick or concrete; numerous coal beds are worked through one shaft; the crosscut rock tunnels or levels, driven to intersect the coal beds, are lined with brick or concrete blocks where needed instead of timber; the coal beds dip more steeply and are not as thick as in most fields of the United States."

Sep 1, 1946

By Michael P. Robinson

It has been suggested that streaming potentials are not nomlally logged because the streaming potentials known to be generated across mud filter cakes are substantially cancelled by streaming potentials generated in contiguous shale beds. Laboratory measuments of streaming poterztials acrons hand specirnens of shale have been adduced as evidence to support the suggestion. The experiments show that streanzing potential depends on the resistivity of the equilibrating solution and on pressure drop. Meizsurenlents were made perpendicular to the bedding planes of the shales. The mechanism of streaming potential generation in shale which the suggested cancellation theory iniplies has been critically examined. Conclusions reached are tested by comparing self potentials logged before and immediately after changing the resistil~ity (but not the weight) of mud in a borehole. The resultant SP change is found to equal the theoretical electrochetnical potential change, i.e., cancellation of .strearning potential is independent of mud resistivity. Since streaming po/entials generateal across mud filter cakes depend critically upon mud reistivity, shale .strearlling potential in situ must he equally sensitive. Using plausible values for the pemleahility, porosiy arzd compressibility of shale, the van Ererdingerl-Hurst approach shows that the pressure drop between horehole pressure and formation pressure in a radially honzogeneous shale is unlikely to coincide with the zone of filtrate invasion. If, perchance. this coincidence occurs with one mud a chatzge 10 a second must result in non-coincidence, i.e., a .YP logged in the second mud should show a streaming potenlial. The dilernrna may he overcome by postrrlating: (I) perrneabilities of shale7 parallel to their beddine planes greatly exceed their cross-bedding permeabilities; (2) a resistivity-sensitive mud cake of low permeability forms either on the face of, or probnhly in the lanlinations of, many shales. These postulates are examined with reference to the pres.sure and SP histories of the McClosky and Aux Vases zones in the Illinois Basin and some srcpport for them found. It is concluded that a definitive solultion to the streaming potential problem does not exist. INTRODUCTION Recently Gondouin and Scala' published a contribution to the streaming potential problem which is notable particularly for the elegance of its experimental techniques. The conclusions drawn by Gondouin and Scala from their studies amplified those previously suggested by Schehck'. In essence they suggested that the streaming potentials known to develop across mud filter cakes" are not normally observed on SP logs because they are largely cancelled by streaming potentials generated in adjacent shales. This suggestion appears so reasonable that many workers in the logging field will be prone to accept it. We incline strongly to the view that the suggestion is basically correct. Nevertheless, it is important to realize that the cancellation hypothesis in the form put forward by Gondouin and Scala does not appear to meet all the known experimental facts. It is the principal intention of this note to call attention to what appears to be a flaw in the relevancy of the experimental evidence adduced by Gondouin and Scala and to suggest an alternative hypothesis to remove it. The consequences of the new hypothesis are compared with field data which, we believe, are unique. A second but not less important purpose of this note is to emphasize that even the new hypothesis does not fully explain all field observations. Hence we conclude that further study of the streaming potential problem is required before it can be said that a definitive solution to it has been attained. The bare experimental bones of laboratory data on streaming potentials appear to be as follows. 1. A streaming potential is developed across a mud filter cake. This streaming potential is virtually independent of cake thickness but depends on the mud resistivity. The relationship is of the form, 2. A streaming potential is developed across a shale. This potential is independent of shale thickness but depends on the resistivity of the permeating solution if the shale has been first equilibrated in this solution. The relationship is of the form, The relationship in (1) was determined first by Wyllie3, confirmed by Gondouin and Scala and qualita-

By D. W. Partridge, W. Lamb, A. E. Lamb, S. Bird

Due to the constant need to improve SAG mill performance and efficiency, extensive research into all aspects of SAG mill optimization has been performed. This paper will review the current evolution of SAG mill shell liner design and their performance at the Copperton Concentrator. The effect of these liner design changes will be discussed with respect to: • Liner configurations • Mill operating performance, i.e., throughput, power draw, mill control • Mill mechanical performance, i.e., liner wear profiles, liner life, liner breakage, liner hardness • Foundry manufacturing considerations • Economic evaluation of throughput versus liner life This paper will discuss the results of Kennecott Utah Copper’s (KUC) and ME International’s (MEI) test work, providing an insight into the past, present and potential future opportunities of SAG mill liner design.

Jan 1, 2002