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Chemicals derived from petroleum raw materials satisfy so much of our industrial needs today that it is good to pause and realise two things -ò the petroleum chemical industry is so new, and the use of chemicals is a very old one.Petroleum is such an'abundant and convenient source of raw material that in little more than 30 years it has become the major source of supplyfor chemical rdanuf ctiree This paperw~l'l 'disc'uss t1.e main 'r`easons' for this rapid development by considering the : sub jest -under tkzree headings ea. Source of Raw Materials.b. The Impact of Petroleum Chemicals on Plant Design and Processes. ce Scme of the Major Chemicals Used Today.

Jan 1, 1962

By M A. Ohsberg, A C. Robertson

Quarrying in Australia accounts for a total annual production of around 200 Mt/a of sand, aggregate and materials used in construction. This material has an annual value of the order of $2 billion and, combined with transport costs of a further $1 billion, the annual value of quarry products are in excess of $3 billion. Currently all of the production of construction materials in Australia is from open cut quarrying or dredging. This paper focuses on the extraction of material by underground mining methods, termed æunderground quarryingÆ. The paper considers a case study for future underground development expansion of the existing open cut Mt Coot-tha quarry, located 5 km from the centre of Brisbane, Queensland.

Jan 1, 1999

By Waddell S

A new procedure has been developed for the direct determination of metals loaded on activated carbon. It involves preparation of a suspension of the powdered carbon in water, followed by direct sampling of the suspension with a micropipette,and analyses by atomic absorption spectroscopy using carbon rod atomization. Advantages of the method include the simplicity of sample preparation and the ability to provide a complete metal analysis in 1 - 2 hours. This enables carbon-in-pulp operators to respond quickly to changes in plant efficiency arising from carbon fouling. The successful application of the method has been demonstrated for gold, copper, nickel and iron. Other elements to be surveyed include calcium, magnesium and silicon.

Jan 1, 1984

In the exploration of an area proposed for the establishment of an underground colliery, consideration of all land and mining titles and consents has early priority. Environmental aspects have become more meaningful. Preliminary assessments of previous geological knowledge, together with new geological and geophysical data from surface investigations, condition the decision to proceed with costly diamond drilling programmes. Methods other than fully cored diamond drilling have a useful application in appropriate circumstances. Drilling exploration is conducted in stages of increasing density of-hole spacing to the point where, with holes at 1 km centres, reserves are categorized as measured. The programme is subject to constant review in the light of market changes.

Jan 1, 1976

Following a period or small-scale activity a joint venture was formed between Ceramco Limited of New Zealand and Georgia Kaolin Company of USA. The initial concept was to process clays for the paper industry but it soon, became apparent that insufficient markets were attainable for the grades of china clay that could be produced for an economic unit. Ceramco Limited acquired the Georgia Kaolin interest in 1975 and proceeded to search for a market, of markets, that would: 1. utilise the unique properties of the product, and 2. enable a viable operation to be created. This paper describes the mineral deposits, the process, the properties of the product, and the marketing methods of New Zealand China Clays.

Jan 1, 1984

By J Ralston, R Sedev, D Fornasiero

The foam behaviour of low molecular weight polypropylene glycols (PPG) was investigated as a function of concentration and molecular weight (190 - 2000 g mol-1). For each polypropylene glycol, foam stability increases with concentration and passes through a maximum, beyond which foamability is suppressed as the solubility limit of the glycol is exceeded and droplets of glycol form. Light scattering data as well as static and dynamic surface tension results provide the key information leading to these interpretations. A maximum in foamability was observed for the PPG molecules with increasing molecular weight (caused by a change in molecular conformation at the interface). This suppresses the Marangoni effect and leads to a decrease in foam stability.

Jan 1, 2005

By Paine G

Kidston Gold Mines commenced mining in 1984 and currently mine approximately 10 million tonnes per annum of ore and waste using a conventional truck and shovel open pit method. The overall mine grade is a low 1.82 grams per tonne.The mineralisation is generally uniform within the ore zones and mining boundaries are staked out from blasthole assays according to pre-determined cut-off grades. Buffer blasting is necessary to minimise post-blast lateral movement of in-situ grade for digging control.The characteristics of the rock currently mined vary widely from soft oxidised to hard unweathered rocks. Blasting problems have been experienced during the initial mining phase in these harder rocks.

Jan 1, 1986

By Campbell D. L, Nutt C. J

The Delta two-degree quadrangle is located in west-central Utah, in the eastern Basin and Range physiographic province but near the intersection of the Basin and Range, Rocky Mountain, and Colorado Plateau provinces (Fig. 1). The quadrangle lies in the region of interior drainage that includes western Utah and most of Nevada, and, like the rest of that province, it has an arid climate. This work is part of the United States Geological Survey's ongoing CUSMAP (Conterminous United States Mineral Assessment Program) project in the Delta quadrangle. Trace-element geochemical data presented here includes information from approximately 600 heavy- mineral-concentrate and 2500 rock sample sites from throughout the Delta quadrangle.

Jan 1, 1988

Remnants of Plio-Pleistocene hot spring systems in Northland, New Zealand contain sulfide minerals, principally pyrite, marcasite and cinnabar. The climate is warm temperate to subtropical, with up to 3 m of rainfall per annum, and rocks are deeply weathered. Decomposition of the iron sulfides in natural outcrops and human excavations results in acid rock drainage with pH typically between 2 and 5, and locally down to pH 1. Deeply weathered basement greywacke is clay-rich and has negligible neutralisation capacity. Some overlying volcanic flows have resisted weathering, and pH is kept near neutral by relict secondary calcite in small areas. Natural rainforest soils commonly have pH between 4 and 5 over mineralised and unmineralised greywacke. Mineralised zones are enriched in arsenic, which occurs in solid solution (up to 1 wt per cent) in sulfides and phosphate minerals. Mercury also occurs in solid solution in some of these minerals, as well as in cinnabar. Base metals are present in low (background) concentrations only. Decomposition of the iron sulfides and phosphates releases the solid solution arsenic and mercury into the acid rock drainage solutions. Arsenic is readily adsorbed on to iron oxyhydroxides formed in acid rock drainage situations, and can be effectively removed from the biosphere. However, this arsenic is released back into the environment if the pH rises, as adsorbed oxyanions become less strongly bound and are displaced by hydroxyl ions. Behaviour of dissolved mercury is similar to that of arsenic but less well understood in these settings. Both arsenic and mercury are readily leached into pH 6 solutions from solid material collected from acid rock drainage sites. These observations suggest that neutralisation of acid rock drainage sites is undesirable in this setting, as arsenic and mercury will become more mobile in the environment. The naturally low pH of 4 to 5 of the forest floor, or even lower pH of natural acid rock drainage, is a preferable management option.

Jan 1, 2003

By R G Warren

Most orebodies have cross-sections of less than 1 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 defined 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 fonn 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 of fluid-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: 1. sources of the mineralising fluids and transporting ligands; 2. sources of the 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 throughout time and space and are mappable on a regional scale. The delineation and empirical prediction of mineral systems can thus be approached by combining two components of metallogenic research. Firstly, the development of maps of regional geoscientific data and data sets of the distribution of factors significant for mineral systems in general. Secondly, the determination of the essential ingredients of particular styles or types of deposits in terms of mappable elements that any exploration program could focus on.

Jan 1, 1994

By Vynne J. F

Drill monitors for rotary blasthole drills have been available to the mining industry since the 1970s. The early monitors were pen strip chart recorders that were developed for the oil drilling industry and adapted to mining. They recorded drilling parameters such as depth, pull-down pressure, torque, RPM, and hailing air pressure. While these recorders could generate a large volume of paper, the amount of useful data was minimal. The data was not summarised and required an engineer to interpret the chart traces. Being a mechanical recorder, the devices were also susceptible to frequent failures. Modem electronics, and the development of the personal computer and its associated software have radically changed the type of instrumentation that is now available for blasthole drills. Perhaps more important is what can be done with the data. The PC allows the logged data to be analysed, summarised, and a variety of reports to be generated. The reports can be customised to meet the specific requirements of the mine and reduce the amount of time that is required to interpret the data. There is also an increased awareness in the mining industry of the importance of production drills. Blasthole drills have typically been the piece of production equipment that received the least amount of attention in a surface mine. Management focussed on the trucks and shovels or draglines. This situation is changing, and therefore the benefits that can be derived from monitoring drill performance are being recognised. Potential economic benefits of installing drill monitors have been documented in previous papers (Peck and Vynne, 1992). Mines are finding that pay-back periods of less than twelve months and rates of return in excess of 30 per cent are not uncommon. These benefits include the following: 1. Reduced overdrilling; 2. Increased rate of penetration; 3. Helping to locate coal seams; 4. Identification of weak and strong zones; 5. Reduce geophysical logging; 6. Optimising blasts; 7. Improved fragmentation and diggability; 8. Decreased coal dilution; 9. Increased bit life; 10. Reduced maintenance costs; 1 I . Improved accuracy of data and reports.

Jan 1, 1995

By Kneisley RO

This paper presents a brief overview of current bump mechanics theories and pillar design methodologies, and relates these concepts to experiences at two mines located in a north- central Utah coalfield where different pillar designs were used to control mountain bumps. Experiences gained at the first mine demonstrated the successful implementation of a two-entry, 9.8 m (32 ft)-wide, yield-pillar design. A U.S. Bureau of Mines field study quantified the timing of chain pillar yielding and resulting load transfer from the gateroad. In-mine pillar response, although apparently sensitive to site-specific conditions, compared favorably to estimates derived using two yield- pillar design methods. A second study conducted at another mine located in the same district, but subjected to different geologic conditions, documents the unsuccessful attempts to employ progressively narrower three-entry pillar designs based on successes achieved at the first mine site. This second mine never achieved a true yield-pillar design. Use of pillar widths ranging between 9.1 m (30 ft) and 27.4 m (90 ft) always resulted in violent bumps in the tailgate pillars. The pillars were either too large to yield, or too small to support peak operational loads. Hypothetical gateroad systems were evaluated using both analytical and empirical approaches for configurations comprised of two rows of conventional pillars, and systems incorporating both yield and abutment pillars. Analysis concluded that yield pillars less than 6.1 m (20 ft) wide and abutment pillars ranging between 30.5 m (100 ft) and 39.6 m (130 ft) square would be required to achieve a stable gateroad design. However, results of a field study conducted on a two-entry, 36.6 m (120 ft)- wide abutment pillar concluded that abutment pressures from the first panel overrode the pillar, and that a still larger pillar may be required to preclude bumps in the tailgate during second panel mining. Without the benefit of a demonstrated in-mine success, it is not clear which design would ensure elimination of gateroad bumps.

Jan 1, 1992

By Sneddon J

This Is not a highly technical paper, but rather an attempt to give a general outline of coal mining for the benefit of those present who may have little or no experience in the industry, and is intended to give some general knowledge which may help to a better understanding of the problems of the industry. A knowledge of the disposition of workable material is necessary to give a comparison with other mining practices. Coal Is usually found in comparatively fiat seams of varying thickness, whereas other minerals are generally found in highly inclined lodes of variable width. However, even in comparatively flat seams unaffected by faulting, grades can, and do, vary greatly over short distances.This variation causes many local problems in working the small area affected. Even very steep seams, and some are much steeper than 1 in 1, vary considerably not only in grade, but also in direction of grade and level course. This is probably caused by folding during alteration of the shape and form of the basins after deposition.These swellies are common to most coal seams and vary only in extent. This type of variation in grade and direction of grade, while common to most ore bodies, is a matter of great importance in coal mining when it is considered that roadways are driven in the seam itself and have to be graded and made suitable for haulage purposes.The area of ground involved in an individual coal mining operation is often very large; many times larger than most metal mining propositions, and often comprises many thousands of acres.When it is considered that the specific gravity of coal may be only one-fifth of that of some ores mined, and the coal seam may be five or six feet in thickness, the vastness of the area necessary for continuous mining operations over many years will be realised.To achieve comparable outputs, with about five times the bulk of material to handle, much more ground and working places must...

Jan 1, 1947

Chile, which hosts around 30 per cent of the worldÆs copper resources, is now the largest producer and exporter of copper. Chilean copper output accounts for more than one-quarter of total worldwide production. An increase to around 40 per cent is forecast by the end of the century. In addition, Chile ranks eighth in silver production and 11th in gold output. In Chile, mining generates more than eight per cent of the countryÆs GDP and the total value of mineral exports is more than US $7.85 billion corresponding to more than 25 per cent of South AmericaÆs mineral production. The rate of growth of Chilean mineral production is indeed impressive. Table 1 shows that in the period 1969 - 1998, copper annual production increased 564 per cent from 667 000 t to 3 764 000 t. Gold production went up 27 times from 1827 to 49 770 kg. Silver output grew 1219 per cent from 95 650 kg produced in 1969 to 1 166 380 kg in 1998. Several facts explain this dramatic increase of production, but the most important was the success of exploration sustained over the same period. This resulted in the discovery of several world-class deposits which when developed into mines, account for much of the production increases. Therefore, this summary review shows the success of basic exploration carried out during the last 30 years in the Chilean territory. The review is based on data already published in different specialised mining books or magazines. It only includes those world-class discoveries not the lesser discoveries important at local scale. All figures were calculated using 1998 $US dollar and average estimated metal prices valid for 1999: copper $US 1653 per t ($US 0.75 per pound), gold $US 9452 per kg ($US 294 per ounce) and silver $US 177 per kg ($US 5.5 per ounce).

Jan 1, 1999

By C Kusi-Manu, E Nyamusika

Geita Gold Mining Limited (GGML) is owned and jointly managed by Ashanti Goldfields Company (Ghana) and Anglogold (South Africa). The mine is currently one of the largest gold producing open pit operations in Africa producing 620 000 ounces of gold and mining in excess of 70 000 000 tonnes of material per annum. The mine currently has 2000 full time employeeÆs and a mine life in excess of 12 years. Geita Gold Mine is comprised of 11 different gold deposits situated in the Lake Victoria Goldfields of North Western Tanzania (Figure 1). The gold deposits are primarily associated with banded ironstone formations within an Archaean greenstone belt. Current global resources, including Inferred, stand at 129.5 million tonnes grading 3.91 g/t (16.2 M oz). Open pit reserves total 70.4 million tonnes grading 4.16 g/t (9.4 M oz). The majority of the gold resource is hosted by the three deposits (Nyankanga, Lone Cone and Geita Hill) within the æGeita TrendÆ a large and continuous (>4 km long) mineralised zone (Adofo, 2002). Geita Gold Mine commenced operations in June 2000, as the second large-scale open pit gold mine in Tanzania. The new workforce of 800 people had no previous experience in open pit gold mining. Since the start of the operation, Geita Gold MineÆs management ambition was to become a world leader in the application of best practice in all technical disciplines and in operational success. Staff development and training was identified as one of the major components in the company achieving its aim. Geita Gold MineÆs management has committed the necessary training staff, resources and the commitment of over $US 1 M a year to staff training and development programs for 2003 and 2004. The process and management of Geita Gold MineÆs mining geology staff training and development program is presented in this paper.

Jan 1, 2003

By P R. G Milne

A risk-based approach to optimising pit wall slope angles was trialled in the design of a 60m deep open pit in extremely weathered and distinctly weathered materials in the Kalgoorlie area. The pit used for this approach was Stage 1 of a three-stage open pit project. The Stage 3 pit will require long term stability, as it will contain the portal for a long life underground project. Information from failures induced by steeper wall angles used in the Stage 1 pit would be valuable in predicting and preventing long-term stability issues in Stage 3. In this approach, the pit wall slopes are based not on the normal factors of safety, but on the assessment of the probability of failure and the direct and indirect costs of dealing with the failure when it occurs. In this example, the benefits and risks of adopting steeper wall angles were compared in a series of graphs. These charted the increasing probability of failure and simultaneous decreasing factors of safety with increasing pit wall slope, and the range in the possible volumes of circular failures for increasing wall angles. These were brought together in a graph showing the potential excavation savings (in BCM) and the probable cost of a failure (probability of failure times the probable volume of the failure). The resulting pit design was based on the project Mining EngineerÆs semi-quantitative assessment of where and when a higher risk of failure was acceptable and where it was not acceptable as it may have unduly affected the integrity of haul roads or production. The design incorporated wider than normal berm widths to accommodate anticipated failures and to enable modification of the wall configuration if required. Several re-designs were required during the life of the project in response to aggressively induced failures. It is considered that this risk-based approach provides a practical route to optimal pit design and operation but it requires an experienced, dedicated and efficient mining team for it to be implemented with benefit.

Jan 1, 2003

By N S. Rauert

The Barrick Gold Granny Smith underground operation located at the Wallaby deposit is currently in a transition period from a near surface operation in ramp up mode to a fully operational deep mining operation with annual production exceeding 1 Mt/a. A key element in this process has been the project execution of the Wallaby Deeps project from initial scoping study in 2007 to the major production component of the Wallaby operation by 2013.The Wallaby Deeps project serves as an example of how an ongoing detailed study program can be integrated into a day to day mining operation assessing risk and mining application in a perceived challenging narrow vein geotechnical environment prior to mining. A detailed study phase to feasibility level is carried out prior to the mining each zone of Wallaby with mining approval only granted on study completion with full justification.A central element in this process is an optimisation study that utilising the Mineable Shape Optimiser (MSO) and Schedule Optimisation Tool (SOT) software tools developed by AMIRA Ltd PRIMO research project is used to carry out an NPV value based option study. This optimisation process being then tempered to reality by a predefined practicality assessment process utilising extensive peer review by operational and study team members as well as industry experts.Other challenges facing the Wallaby Deeps project include the requirement to maintain the ramp up production profile to full production by 2012 without disruption but with pressures of an ongoing geological resource to reserve conversion program essential for the study process and mining approvals and also development rate achievement.To date these challenges are being met with integrated planning and the value of the study program itself. Production in 2012 year to date is well on the way to annual production rate exceeding 1 Mt/a.CITATION:Rauert, N S, 2012. Wallaby Deeps - from the optimal opportunity to reality, in Proceedings Narrow Vein Mining 2012 , pp 207-214 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 26, 2012

ON milling plants, where a large number of slime samples are collected each day, considerable work is involved preparing them for the assayer. The usual practice (as the material usually passes a 150-mesh screen) is to place the sample on a 30-mesh screen an 1 brush it through on to a cloth below. The sample on the cloth is mixed and a portion taken out for the assayer. A simple apparatus that would rapidly disintegrate and prepare these slime samples for the assayer was needed. The manufacturers of assayers' supplies appear to have overlooked apparatus for this class of work. To overcome the tedious preparation of slime samples by the above method, the writer devised the apparatus shown in the accompanying drawing (Fig. 1). It will be seen that the apparatus is a combination of a modified household flour sifter and a riffle sampler.The apparatus is made up in two parts, the top half consisting of a screen box carrying the disintegrator, and the bottom half a, riffle divider, the joints between the two being a slip-in fit. A lid is provided for the screen box. The disintegrator consists of an open drum, to which are soldered three rows of spring steel fingers. Each row has a number of steel fingers which, on rotation, press against the semi-circular screen in the screen box, forcing the slime particles through.To operate the apparatus an 8-in. sample tin is placed under each spout of the divider. The sample of slime, which is about 2 lb. in weight and usually in lumps, is placed in the screen box, the cover replaced, and the drum rotated about half-a-dozen times by means of the handle.The disintegrated sample is divided by the rifle divider into two portions, which are found in the tins under the spouts of the apparatus. One of these samples is taken, mixed on a sheet, and sent to the assayer.

Jan 1, 1922

By S Prakash, P S. R Reddy, S K. Biswal, V N

The beneficiation potentials of two Indian iron ore slime samples by flotation technique using both cationic and anionic collectors are investigated. The direct and reverse flotation of slimes using oleic acid and dodecylamine as the collector have been studied using both conventional cell and flotation column to upgrade the iron content. The characterisation studies of iron ore slimes indicated that the samples are extremely fine in nature and contain substantial amounts of martite, magnetite, goethite, limonite, kaolinite and quartz beside hematite as the dominant iron mineral phase. The samples on average contain around 52 - 58 per cent Fe, four to seven per cent SiO2 and five to six per cent Al2O3. Taking into account the physico-chemical properties of the slimes, it is attributed that both the samples need to be deslimed by hydrocyclone before flotation. The results of both cell and column studies on sample I showed that a feed containing 57.5 per cent Fe could be upgraded to 65.5 per cent Fe at 50 per cent overall recovery with simultaneous decrease in SiO2 and Al2O3 content to 2.3 per cent and 1.8 per cent respectively by using oleic acid as the collector. In the case of sample II, it has been established that reverse flotation of silica is the better technique to improve both the grade and recovery of iron. Batch flotation studies in cell indicated that a product with 63.1 per cent Fe could be obtained at 40.0 per cent yield with 63.6 per cent iron recovery by the combination of reagents such as gum acacia, sodium hexa meta phosphate and dodecylamine. The performance of column flotation techniques is far superior to that of the cell with respect to both grade and recovery. The concentrates are found to be quite suitable for commercial exploitation as pellet feed.

Jan 1, 2005

By A Xenidis, N Papassiopi, I Paspaliari

Acid generating sulfidic tailings covering a surface area of approximately ten hectares have been deposited in the historic mining area of Lavrion in Greece. For the development of a remediation strategy, the performance of three alternative acid generation control techniques was assessed under actual field conditions. The techniques involved the application of dry covers, including a synthetic HDPE liner and a compacted clay layer with capillary barriers. Furthermore, the potential formation of a low permeability hardpan layer at the sulfidic tailings surface by limestone addition was investigated. The field tests covered an area of 2500 m2, which was divided in equal quadrants, corresponding to the control plot and the three alternative remedial options. To monitor the amount and quality of percolated water, lysimeters were installed beneath the four plot areas. Based on the monitoring results over a period of six years, due to the dry climate encountered in the Lavrion area, a low amount of water percolates through the pyrites if no action is taken (control area), corresponding to 20 - 25 mm per year. This water is highly acidic, pH = 1.2, and contains a heavy load of contaminants, eg Fe 40 g/L, Zn 15 g/L, and As 1.5 g/L. The HDPE cover eliminated the percolation of water into the material for a period of five years, whereas thereafter the volume of leachates increased to values similar to those recorded in the control plot. The water collected in the lysimeter of the clay cover has been minimal corresponding to approximately five per cent of the drainage volume produced in the control area. In the fourth plot, which involved limestone addition to sulfidic tailings, the volume of drainage initially collected was almost double compared to the control area. After three years of monitoring, significant reduction in the amount of percolated water was observed, attributed to the formation of a cemented layer (hardpan), at a depth of 40-80 cm from the testpad surface.

Jan 1, 2003