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As an indication of the overall quality for the three periods the expected run of mine ashes are given. Operating Period R.O.M. Quality Years 1 - 4 22.0% Ash (db) Years 5 - 14 18.8% Ash (db) Years 15 - 20 17.3% Ash (db) The size distribution of the coal was found to be generally as predicted in the 1980 study. However the ash, particularly in the smaller fractions, was found to be higher due to the previously mentioned friable nature of contaminant material. Following evaluation of this additional information, the proposed washery design was reconsidered. Three basic process schemes were examined. 1. A jig with two-stage heavy medium cyclones and spirals for treatment of three different size ranges. 2. A heavy medium vessel, two-stage heavy medium cyclones and spirals for the different size fractions. 3. A jig for initial treatment of the total feed (that is all sizes) with single-stage heavy medium cyclones and spirals for rewashing of the intermediate and smaller size Fractions.

Jan 1, 1986

In line with the recommendations of The AusIMM/AMIC Joint Committee Report in 1981 Isa Mine ore reserves are quoted in proven and probable categories (recoverable ore reserves). Internal reporting considers geological mining and extraction reserves in terms of proven and probable categories. Proven reserves are subdivided into detailed and confirmed ore. External reporting includes a possible ore category (not a reserve). The basis for these divisions is the spatial distribution of sampling data. Proven ore is defined by data 80m apart or less and probable ore by data 80m to 150m apart. Within the proven ore category detailed ore is defined by data 40m apart or less and confirmed ore by data 40m to 80m apart. Kriging is used to provide tonnage and grade information for the currently producing 1100 and 1900 orebodies. Kriged cross sections are initially produced to coincide with existing geological sections and are used for stope design. Stope ore reserves calculations are ideally performed at three discrete design stages, which in order of increasing accuracy are: preliminary design (PD), final design (FD) and ring design (RD). At each of these stages Kriged meshes incorporating the most recent sampling information are produced for reserve calculations. Historically derived mining and extraction factors are applied both to global reserves and individual stope reserves to reflect changes in geological and mining conditions as well as the sequencing of extraction. These factorised reserves are then inserted into the computerised reserves inventory system (RIN). Monitoring of individual stope performance continues into the production stage with input from planning, research, survey and production personnel. This allows actual extraction factors to be derived for comparison with predicted factors, providing some measure of stope performance.

Jan 1, 1987

The scope of this paper has been restricted to the manufacture in Australia of seamless steel tubing, as the subject of buttweld tubing has already been dealt with by A. N. Hamilton in a paper delivered before the Institute in 1935.The earliest attempts to produce seamless steel tubes were based on the extrusion of lead piping, a process developed by Thomas Burr in 1820. However, the problems associated with the extrusion of white hot steel had to await the metallurgical knowledge of the twentieth century, and these early attempts were unsuccessful.The years 1840 and 1845 were noteworthy for the introduction of the first practical methods of making seamless steel tubes. In one process, a large circular plate was 'cupped' in a vertical press; in the other, a round steel billet was pierced for its entire length.The product from both these methods was then elongated by repeated reheating and rolling over an internal mandrel, in a circular pass mill.At this juncture, it should be noted that every process developed for the manufacture of seamless steel tubing can be divided into two parts. In the first part of the process, attention is centred on the conversion of the heated bloom into a short, thick-walled, hollow cylindrical forging; this product is then elongated in the second section of the process. It will be noted that these two sections are actually quite distinct, and in practice numerous methods exist for carrying out both functions. The first stage is generally referred to as 'piercing', the second stage being termed 'elongating.' In this paper, the author proposes to outline the various methods adopted for piercing and elongating; the reader is asked to note that although certain methods of piercing are almost invariably associated with only one method of elongating, the distinction between the two functions is, nevertheless, a vital one and should be noted accordingly.

Jan 1, 1947

By Chung S

The problems associated with equating calculated and measured properties of explosives with blast results are examined. The results from six full-scale test blasts in a limestone quarry are described. The purpose of the full-scale trial was two-fold, i) assess the role of varying aluminum levels (10% maximum) in a given explosive in terms of blasting performance, and, ii) find correlation between calculated, as well as measured properties of explosives and the resulting blasting performance. Each blast consisted of 36 holes, initiated with short period detonators in a staggered 3-row 2.4 m x 2.7 m (burden x spacing) pattern. Each 75 mm diameter borehole with no subgrade contained on the average 22 kg of 65 mm diameter cartridged explosive. The bench height was 6.3 m, with an average collpr of 1.2 m. The calculated and mgasured properties of explosive included, strength, detonation and explosion pressure, VOD, plate-dent depth, underwater bubble and shock energy, and the total energy efficiency. Each blast was instrumented to measure detonation velocity, initiation sequence, muck pile motion by means of high-speed photography and blasting vibrations. The post-blast measurements included photogrammetric survey of muck pile for size distribution on surface, muck pile profile, as well as the extent of backbreak. The results showed that the overall blasting performance of the explosive in question was nearly commensurate with the level of aluminum in the explosive, and there was excellent correlation between the underwater test properties and the blast results. The study also exemplified the need for more extensive post-blast analysis than is usual practice, for quantitative assessment of blasting performance of explosives and providing the necessary data-base for calibration of blasting models.

Jan 1, 1990

Caving is the lowest cost underground mining method, provided that drawpoint spacing, drawpoint size and ore handling facilities are designed to suit the caved material and that the drawpoint horizon can be maintained for the draw life. In the near future several open pit mines that produce in excess of 50 000 tons per day will have to examine the feasibility of converting to low cost, large scale underground operations. Several other large scale, low grade underground operations will experience major changes in their mining environments as large dropdowns are implemented. These changes demand a more realistic approach to mine planning than has occurred in the past, where existing operations have been projected to increased depths with little consideration of the change in mining environment that will occur. As economics force the consideration of underground mining of large, competent orebodies by low cost methods, the role of cave mining will have to be defined. In the past caving has generally been considered for rock masses that cave and fragment readily. The ability to define cavability and fragmentation, the availability of large, robust LHDs, a better understanding of draw control requirements, improved drilling equipment for secondary blasting and reliable cost data have shown that competent orebodies with coarse fragmentation can be exploited by cave mining at a much lower cost than with drill and blast Methods.

Jan 1, 1995

By S Arthur, J Mills, A R. Penney

A key aspect of seismic management for narrow vein stoping of the west zone at the Beaconsfield Gold Mine has been the development of a fully remote extraction method that does not require personnel access to sill drives within the orebody once stoping commences. Initially, sill driving within the reef is undertaken in a conventional fashion under geological control. Sill drives are developed at a vertical backs-to-floor height ranging from 7 - 12 m. This mining stage includes all activities prior to the commencement of stoping, after which no personnel enter the unstoped section of the sill drive at any time for any purpose. The key aspect of the remote mining cycle is the use of footwall drill drives paralleling the ore drive at a minimum separation of 6 m under survey control. These drives provide the platform for fan drilling of the orebody. Up to 20 m strike panels are blasted in a single shot using electronic blasting techniques. Ore is mucked from the extraction level sill drive using tele-remote loaders. Aerials for tele-remote operation are installed through boreholes from the footwall drive, negating the need for personnel access to the sill drive. Once mucking of a stoping panel is complete, a laser scanning survey instrument is used to survey the final void for subsequent design and production reconciliation purposes. The survey instrument is mounted in the bucket of the tele-remote loader, and is also operated remotely. Finally, tele-remote loaders are used to dump waste rock fill from the upper sill drive. A system of flags and cameras installed in boreholes from the footwall drive assists the operator in locating the live stope edge during the filling cycle, again negating the need for personnel access.

Jan 1, 2008

Prior to 1975, most geochemical drainage surveys for gold were based on analyses for pathfinder elements, such as As and Cu, rather than direct analysis for gold, for two reasons. Firstly, analytical methods lacked the required sensitivity and precision to determine gold abundance at background levels in stream sediments (usually in the range 1 - 10 ppb). Direct analysis was also inhibited by concern about the potential for gross sampling error, brought about by the dispersion of gold as rare discrete particles of the extremely dense native metal. This sampling error, commonly termed the 'nugget effect' has two components: erratic distribution of the element within the stream environment, related to hydraulic factors, and the difficulty in taking representative samples, both in the field and later in sub-sampling for analysis. In more recent years the collection of very large samples was commonly advocated to overcome the sampling problem (Hams, 1982; Nichol, 1986).Two methods have been widely used to evaluate the gold content of bulk samples (Elliott and Towsey, 1989). The preparation and analysis of heavy mineral concentrates has in recent years been largely superseded by the leaching of gold contained in bulk samples by cyanide, commonly termed the BLEG (Bulk Leach Extractable Gold) or BCL (Bulk Cyanide Leach) methods. The terms BLEG and BCL are essentially interchangeable and the latter will be adopted for the remainder of this paper. The strengths and weaknesses of these two techniques were compared with conventional total gold analysis on fine fraction (

Jan 1, 1976

A pilot plant has been constructed for the low temperature carboni-zation of Wallarah (N,S,W6) coal at the rate of 300 pounds per hour. The plant was built and tested at Australian Mineral Development Laboratories for Broken Hill Pty, Ltd. Char product was required within the specifi-cation 15.5 `S' 1 per cent volatile matter. The design was based c atz an entrained state carbonizer developed by Parry and U.S. Bureau of Mines co-workers at Denver, U.S.A. The method involved the carbonizatioan of fine coal by entrainment with air in an externally heated reactor.. The heat required for carbonization of tits coal was supplied to the be, both through the reactor walls and also by internal cambustion of products .

Jan 1, 1963

By E Rubio, N Morales

Currently, mine plans are optimised by using several criteria as objective functions, like profit, life-of-mine, concentration of some pollutants, mining costs, confidence level or ore resources, with consideration of constraints related to production rates, plant capacities and grades. Whilst this approach is successful in terms of producing high value production schedules, it uses a static opening sequence of the drawpoints and therefore the optimisation is made within the level of freedom left by the original opening schedule, and as a result it is far from the true optimal value of the project. This paper presents a model to optimise the sequence of drawpoint opening over a given time horizon. The emphasis is in the precedence constraints that are required to produce meaningful operational sequences considering the exploitation method (panel or block caving), physics considerations and logical rules. Further on, while it applies the standard approach of maximising net present value (NPV), it considers other targets for optimisation, like the robustness and constructability of the plans. Finally, it applies the model to real data to obtain feasible plans by means of this model. The first tests were done varying the production capacity of the mine.

Sep 26, 2011

By Turco E, Pelizza S, Castellani F, Stanganelli V

Excavation are in progress for the building of tunnels of the waterworks in West Campania (Naples). The works afferent to part II envisage tunnelling through a miocenic clay formation over consolidated by two tunnels whose sections are respectively 50 and 10m2. Tunnel blasting is carried out by mechanical means. An excavator is used in the big tunnel, whereas in the small tunnel a roadheader machine is used for a short tract and a mechanized shield is used over a long stretch. The tunnel feature complex problems regarding stability. This is particularly so for the big tunnel whose face appears unstable. For this reason the clay on the face is sistematically bolted by means of 3-metre and 12-metre bolts. This prevent the use of a TBM for tunnel Excavation.

Jan 1, 1987

By P Ireland, S Mitra, G Evans, A Wang, M M. Hoque

This work presents an experimental study of flow dynamics in a bubble column system (crosssection: 100 mm × 100 mm) using a non-intrusive optical technique with targeted application in a novel REFLUX™ Flotation Cell. The study had two aims – measurement of the bubble size distribution (BSD) and local energy dissipation rate in the system due to bubble motions over a low range of gas flux varied from 0.02 to 0.08 cm/s without impairing the visibility of the system. A central bubble plume was generated by injecting compressed air through a sintered glass distributor in quiescent liquid. Highspeed shadowgraphy imaging was performed to determine the BSD in the presence of surfactant (Sodium Dodecyl Sulfate (SDS) concentration ~ 10 ppm). The BSD exhibited a unimodal shape at different locations of the column for all gas flux cases. A non-intrusive particle image velocimetry (PIV) technique was used to quantify the flow field generated by the bubble motions, which was used to estimate the local energy dissipation rate. The local energy dissipation rate exhibited an increasing power-law dependency on the gas flux.

Aug 24, 2022

By Lynce AJ

A method has been developed for the prediction 'of 'the- . performance of a multi-stage grinding=cyclone classification circuit of any degree of complexity on any ore for a wide range `of operating conditions, The-.method involves ;the. s` inu ation b'f the circuit on a digital computer by means of. matheinatical' models ":bf the"' individual units:, the. development of which is- also discussed'., At present, in order to establish 'the!'" charac'teristtics of they ¦ore- and the machines. in mathematical terms;. it is necessary to know'the performance of a , simpl,.e. ; system on tri.e ore - for! one known ,,feed rate and.fe.ed s,l-zing, analysis. ' ,Fiiither work on :the- separate identification 'of the terms defining the characteristics of the ore and the machines may permit the method to be, applied directly to the cj.4ta,derived from small scale tests on ores,

Jan 1, 1964

By D Macklin, R Hammond, T G. Blenkinsop

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. Base metal sulfides at the Currawong volcanic-hosted massive sulfide deposit in the Lachlan fold belt of Victoria, south-east Australia, were formed in the late Silurian, but D2 deformation of the deposit in the Devonian Bindian orogeny was responsible for structural repetition that created the resource. The primary sulfide lenses were stacked by reverse-sense shear zones within a volcano-sedimentary sequence dominated by dacitic intrusions and turbidites. Shear zones were localised along lithological contacts, including the margins of the sulfide lenses. The competent dacitic intrusions did not fold during D2; folding was restricted to the packages of metasedimentary rocks. Overturning occurs within, but not between, the metasedimentary packages. The overall geometry of the deposit is that of a duplex, but the detailed shape does not conform to idealised duplex types, probably because of post-D2 deformation. A significant complicating factor for understanding the deposit, and for resource evaluation, is the dramatic variation in deposit geometry along strike, which may also be due to post-D2 deformation. The present day configuration of the deposit results from constructive thrust repetition of what may have been a single ore lens and destructive along-strike disruption controlled by rheology.CITATION:Blenkinsop, T G, Macklin, D and Hammond, R, 2015. Rheological controls on the geometry of the Currawong volcanic-hosted massive sulfide deposit, Lachlan fold belt, Victoria, south-east Australia, in Proceedings PACRIM 2015 Congress, pp 469–474 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

By E E. Hall, C D. Taylor, D S. Yantek

Regular monitoring for methane gas is required near working faces in gassy underground mines where the potential for methane ignitions is greatest. However, high concentrations of methane can also accumulate outby the face where methane levels are monitored less frequently. Wearing a personal methane monitor equipped with an alarm could protect persons working in these outby areas from injury or death due to methane ignitions. The National Institute for Occupational Safety and Health (NIOSH) has conducted studies to evaluate methanometers that could be used as person-wearable monitors. For the present study, seven different person-wearable monitors that are not currently approved for underground use were evaluated. This paper describes test procedures used to evaluate performance of methane monitors that could be used underground for continuous personal monitoring of methane. The information in this report provides data that allows the reader to make side-by side comparisons of these instruments.

Jan 1, 2005

By Partridge AC

In order to provide a'valid description of the performance of size classifying equipment it is necessary that a standard procedure be established for testing and for the subsequent expression of results. Such a standard would permit the valid comparison of alternative equipment and may provide a basis for the prediction of performance. An Australian Standard (AS 1634-1974) has been used as the starting point for an International Standards Organisation-(ISO) working group (of which the authors are members) charged with the preparation of an ISO Standard for the expression and presentation of the results of tests on coal size classifying equipment. The approach taken is applicable to all types of sizing equipment (cyclones, screens, dedusters, etc.) and should be suitable for adoption in the minerals industry as well as by coal preparation engineers. The background to the proposed Standard, is given,including a critical review of other frequently quoted performance measures, the procedures described in the draft are discussed, and an example given.

Jan 1, 1984

Longwall mining has historically been considered an application for only deep reserves. This has changed in the recent past as longwall has become competitive with surface mining. Many resources marked for stripping now he in the domain of longwall. Today, the best longwall mines compete with strip mines operating in even the lowest stripping ratios. Longwall mining has become more widely applicable than it ever has been in the past and it is now generally considered the method of choice in situations where high levels of coal production are required. But the best way to use longwall mining has become less certain with the opportunities that highwall access provides. Wall-to-wall mining is the use of retreating longwall developed from and retreating toward a highwall. As a surface mine reaches its ultimate highwall, the opportunity for longwall punch mining or wall-to-wall mining is obvious. There are several clear operating advantages to operating a longwall in this way and where highwalls are available, the best economic decision will generally favor planning a direct approach. Yet wall-to-wall is not applicable in every situation even when highwalls are available. Moreover, the advantages of wall-to-wall mining do not always justify the cost of trenching or box cuts to access coal . However, many tactical advantages of operating from the highwall can be realized with the use of specially equipped blind drilled or raised bored shafts. This paper discusses novel techniques for a shaft assisted mining system which permits the addition of conveyor structure and mine systems from the surface to the face for advance and retreat mining with longwall, shortwall or room and pillar panels.

Jan 1, 1998

Stressed steel cables, similar to those used in pre-stressed concrete, have been used for rock mass reinforcement, in civil engineering, for many years. This type of reinforcement has been attempted for open pit wall support, on several occasions, but has never gained wide acceptance as a practical approach because of high cost and complexity of installation in a mining environment. On the other hand, untensioned, fully grouted cable dowels (cable bolts) have gained wide acceptance for ground support in underground mining and it is perhaps surprising that this technique has not been used to a greater degree in open pit mining. The first systematic application of this technique in Australia was at the Mary Kathleen uranium mine in 1976, where it was spectacularly successful in maintaining steep stable walls in a jointed rock mass with unfavourably dipping structures. Cable bolting, and other techniques, are now being used to an increasing degree for wall support in both small and large open pits, for two basic applications: 1. stabilisation of potentially unstable sections of a wall, or 2. to allow steeper walls and reduce waste stripping volumes. In virtually every situation, wall rein- forcement has significant economic as well as technical advantages. The paper describes the development of the techniques, the mechanisms of wall support, the methods used, and discusses several case histories of its application to Australian open pit mines.

Jan 1, 1986

It is now 32 years since Walsh's classic publication (1) heralded the beginning of atomic absorption spectrophotometry (AA) as we know it today. During this time, AA has developed into the most widely used analytical techniques for the determination of the metallic elements. The years have seen progressive changes in instrumentation so that today it is a highly sophisticated technique. One area of development in AA that is experiencing renewed attention is in the area of electrothermal excitation. In this technique a flame is not used by the analyte solution is atomised by an electrically heated graphite furnace. For a large number of elements the graphite furnace atomic absorption spectrophotometer (GFAA) is considerably more sensitive than flame work. It is common to be able to detect concentrations one hundred times smaller than those possible by flame. Today's geochemists are finding that they can gain valuable information out of very low limits of detection. Electrothermal atomic absorption spectrophotometry (or GFAA) allows for low limits of detection and can be seen to be a valuable aid to geochemical prospecting Analysis.

Jan 1, 1987

Different mineral structures can act as mineral reservoirs for the immobilisation of toxic compounds. In the paper different mineral structures will be explained and their potential for incorporation of heavy metals and anions will be shown. The following structures will be explained (P÷llmann, 2007):apatites (incorporation of heavy metals, silicate apatites);ettringites (fixation of oxo-anions);layered double hydroxides (fixation of heavy metals, fixation of anions, fixation of organics, sorption materials); andmanganese oxides. Different possibilities of solid solution and incorporation in crystal structures will be explained and advantages of stable minerals be proved using experiences with real toxic wastes. In the process of thermal treatment the formation of primary mineral reservoirs will be shown. The precursor phases (similar to well known cement minerals) react in contact with water to form secondary mineral reservoirs (cement hydration mineral products). The formation condition of primary and secondary mineral reservoirs and their stabilities and incorporation possibilities will be shown. Examples for incorporation of various elements are given. Application for water clarification will be outlined.

Jan 1, 2008

Reconciliation of the ore reserve has been an integral part of mine planning since Bougainville Copper Ltd. commenced mining operations at Panguna in 1972. Confidence in the ore reserve is considered essential, both for long term planning of the deposit, and for fair and accurate reporting of the mining resource to shareholders. Recoverable, proven reserves are published annually. Although the ore reserve quoted is based only on diamond drill hole (DDH) information on a 122 metre by 122 metre grid, predictions are regularly reviewed against production. Concentrator results are taken to be the best estimate of grade and tonnage. Surveyed mined perimeters and average blasthole grades are reconciled at the end of each month against the concentrator figures. Predictions of the in-pit reserve are made using an orebody model. DDH and blasthole (BH) sample grades are used to make predictions for both long and short term planning. This paper describes how reconciliations are used to validate ore reserves at Panguna. Development of the orebody model is described, and some of the difficulties encountered in doing reconciliations are addressed.

Jan 1, 1987