Tons Of Ore Research Articles

A stochastic mixed integer linear programming framework for oil sands mine planning and waste management in the presence of grade uncertainty

ABSTRACT The primary purpose of oil sands mine planning and waste management is to provide ore from the mine pit to the processing plant while containing the tailings in an efficient manner in-pit. Incorporating waste management in the mine plan is essential to maximize the economic potential of the mineral reserve and minimize waste management costs. However, spatial variability such as grade uncertainty results in ore tonnage variations, which leads to fluctuations in the quantity of ore to be processed and waste to be managed. This paper investigates the application of a stochastic mixed integer linear programming (SMILP) on oil sands mine planning to integrate bitumen grade uncertainty and waste management. Sequential Gaussian simulation is employed to quantitatively model the spatial variability of bitumen grade in the oil sands deposit. Multiple simulated orebody models are used as inputs for the SMILP model to generate optimal mine plans in the presence of grade uncertainty. The results demonstrate that the SMILP schedule generates 14% and 17% improvements in net present value compared to the E-type and ordinary kriging schedules, respectively. These results indicate that the SMILP model is a robust tool for optimizing stochastic integrated oil sands production schedules and waste management.

Electroosmotic Dewatering of Iron Ore Tailings: A Laboratory Study to Improve Geotechnical Properties

Minerals are subjected to ore processing to turn them into usable and saleable raw materials. First of all, the ore is reduced to the smallest size with the crushing‐grinding process, and after using water and chemical additives, according to the characteristics of the ore, useful minerals are taken, and unwanted minerals are stored in the tailings pools. Approximately, 26 million tons of mineral wastes were generated annually from ore processing facilities in Turkey. The construction of tailings pools, the stability of the tailing sludge, and the safety of dams are a burdensome issue faced by miners. In the ore plants, an average of 3 tons of water is used to enrich 1 ton of ore, and most of the wastewater cannot be removed by traditional methods, causing various economic, environmental, and stability problems. In this study, the dewatering of an iron ore tailing by applying different voltages by the electroosmosis method was investigated in the laboratory environment. By discharging the water of the iron ore tailings by the electroosmosis method, the solid content was increased from 43.01% to 87.63%. Thus, it has been observed that there will be a significant improvement in the geotechnical properties of the tailings’ material. It has been estimated that electroosmotic dewatering rises with increasing the voltage gradient and the energy consumption varies in the range of 0.588–30.645 kWh/dry ton. The void ratio decreased from 5.58% to 0.23%. In the dewatering experiments, different parameters such as the amount of water discharged, density, void ratio, water content, and power consumed were measured or calculated and the relationships between them were discussed with graphics. In electroosmosis experiments, it has been observed that besides the voltage applied in the discharge of water, the mineralogy of the tailings has a significant effect. Since there are serious abrasions on the electrodes used in the experiments, alternative electrodes should be tried.

EFFECTS OF A PILLARLESS, CENTER-OUT STOPING PATTERN ON HAULAGE DRIFT PERFORMANCE AND ORE TONNAGE AT RISK

With respect to mining sequence, this article intends to investigate the impact of pillarless centre-out stoping patterns (e.g. pyramidal sequences) on the performance of mine haulage drifts (e.g. ore access units), the tonnage of unmined ore at risk, and the required quantity of fill material. Using RS2D software, a two-dimensional, elasto-plastic finite-element model for a haulage drift located at 1200m below the surface in the orebody’s footwall has been built. The spread of yielding zones into the rock mass around an access drift and unmined stopes is used to assess mine haulage drift stability and estimate the amount of unmined ore at risk owing to local mining activity. The findings are presented and discussed in terms of the size of failure zones, the number of tonnes of unmined blocks at risk, and the amount of backfill materials required, all in relation to the mining stage. The findings show that haulage drift stability is rapidly diminishing. The drift roof begins to deteriorate at an early stage (after mining step 3). In the drift roof, left wall, floor, and right wall, failure zones measured 1.55m (step 3), 2.28m (step 4), 2.57m (step 5) and 1.88m (step 5) accordingly. After mining step 4, there was a total of 905 m3 of unmined ore at risk (4100 tons), and after mining step 5, a total of 1500 m3 (30 tons) of back fill material was required to strengthen stopes.

The Recovery of Cu, Co, Zn, and Mn from a Complex Oxide Ore Using an Enhanced Reduction Leaching

The processing of Cu, Co, and Zn at the Boleo project in Mexico involves two-stage (oxidation–reduction) leaching to extract a total of 85–88% Cu in 4 h. The first stage is an oxidation leaching using sulphuric acid (120 kg/tonne ore) at an Eh of 900 mV for 2 h. Then, the reduction stage takes place in 2 h with SO2 gas sparging for Mn and Co extraction at an Eh of 350–370 mV. The final extraction rates of metal values are 92% of Mn, 80% of Co, and 60% of Co, respectively, after 4 h of leaching at 70 °C. However, the same metal recoveries were obtained within 2 h using an equal amount of sulphuric acid and the addition of 25 kg of SO2 per tonne of ore in a single stage leaching in this research. In this case, the Fe extracted from the ore as Fe2+/Fe3+ is believed to have acted as an electrochemical couple contiguously leaching the Cu sulphide and Mn oxides, which also increased the Cu recovery as the Cu mineralised mostly intergrowths in these mineral structure matrices. A significant improvement was made in which the leaching time was halved to 2 h compared to 4 h in the previous plant design and current operation, involving the two-stage oxidation–reduction leaching.

Talvivaara Sotkamo Mine – Bioleaching of a polymetallic nickel ore in subarctic climate

The main activity of the Talvivaara Mining Company Plc. is the development and exploitation ofthe Talvivaara deposits in Sotkamo, Finland using bioheapleaching. The Talvivaara deposits comprise one of the largest known sulphide nickel resources in Europe with 1004 million tonnes of ore, sufficient to support anticipated production for a minimum of 45 years. The mine started in late 2008 and will have an annual nickel output of approximately 50,000 tons when it reaches full production. In addition, the mine will also produce zinc (approximately 90,000 tpa), copper (approximately 15,000 tpa) and cobalt (approximately 1,800 tpa) as by-products of the process. The viability of bioheapleaching technology for the extraction of nickel has been demonstrated in a large on-site pilot trial using Talvivaara ore. The three year pilot has shown that the leaching process also works well in the subarctic climatic conditions of Eastern Finland.

Mineral carbonation with thermally activated serpentine; the implication of serpentine preheating temperature and heat integration

In this paper, heat integration was coupled with serpentine preheating temperature optimization and was for the first time applied to a mineral carbonation process. For this, a process was selected with the aim to minimise its heat demand. Aspen® Energy Analyzer software was used for heat integration with pinch analysis. The mineral carbonation plant considered here processes 100 tonnes of serpentine per hour, with Mg content of 237 kg per tonne of rock, corresponding to 0.88 million tonnes per year to treat 0.5 million tonnes of CO2 emitted from a cement plant. In the base case considered for heat integration, 50% of the serpentine magnesium content was converted to hydromagnesite using 30% of the flue gas CO2 which means that 5.85 tonnes of rocks were required to capture 1 tonne of CO2.The serpentine preheating temperature, was optimized as 400 °C. Application of heat integration at this solid preheat temperature reduced the process heat demand by 25% compared to our previous study, as 5.0 GJ per tonne of CO2 captured or 9.5 GJ per tonne of CO2 avoided. This corresponds to 0.86 GJ per tonne of ore as the process heat demand was only attributed to the mineral activation. In addition, the impact of process parameters including the solid-liquid ratio and the dissolution reaction extent on the heat integration strategy was evaluated. It was found that solids concentrations as low as 5% substantially reduced the sequestration efficiency as the process waste heat was not sufficient at the optimized solid preheat temperature. Furthermore, the uncertainty behind unknown parameters such as activated serpentine heat capacity and solid heat transfer coefficient was evaluated to validate the study.

Efficient flotation recovery of lead and zinc from refractory lead-zinc ores under low alkaline conditions

In this study, a new flotation approach, a low-alkaline and non-desliming process, was introduced for improving lead and zinc recoveries, lowering production cost and reducing environmental pollution. Lab-scale experiments results show that the new process contributed to the flotation of the complex mixed sulfide-oxide lead and zinc ore regarding two aspects: (1) High alkaline process (pH = 12±) was replaced by low alkaline process (pH = 9±) by using collector WS (a mixture of ethyl thiocarbamate, ammonium dibutyldithiophosphate and dithiophosphate-25) and combined depressant Na2S/ZnSO4/Na2SO3 for lead sulfide flotation; (2) Non-desliming process was successfully achieved by using collector MA (a mixture of ether amine, hydroxyethyl cellulose and polyacrylic acid) and combined depressant SHP/SS (sodium hexametaphosphate/sodium silicate) for zinc oxide flotation. And 43.37% Pb in the Pb concentrate was recovered, the corresponding Pb grade was 52.73%, total 84.42% Zn was recovered by the flotation of zinc sulfide minerals and zinc oxide minerals. Moreover, the two aspects of the new approach were systematically verified from lab-scale to industrial-scale application. The industrial-scale flotation tests show that Pb recovery in Pb concentrate increased by 1.86% compared with that of original system during industrial-scale tests period, and the Pb recovery increased by 4.09% compared with that of original system before industrial-scale tests period, while the Zn operating recovery in zinc oxide concentrate improved by 19.52%. Moreover, the total reagent cost of the whole new process significantly declined by 3.93 yuan per ton of ore.

Extraction of Sand from Waste dumps of Mining: A New Approach to Address the Environmental Issue of Goa

Goa is India's smallest state by area (3700 sq.km) but rich of mineral deposits such as -iron ore, manganese ore, bauxite and many minor minerals like basalt, laterite stones, rubbles, river sand etc., Exploration and exploitation for iron ore has started at the beginning of the 20th century. Iron ore mining in Goa is completely in the private sector. The mining belts extend over a length of 65km from SE to NW of the State covering around 700 sq.km. The iron ore deposits are distributed over the Northern, southern, and central blocks of Goa. It is the only state with large area under mining. Mineral production in Goa started in the late 1940s. Initially with Manganese ore but later shifted to iron ore with phenomenal growth in production. Every year around 60million tonnes of iron ore was being handled till 2015. Subsequently capping has been imposed by Hon’ble Supreme Court to mine 20 million tonnes from 100 mines by open cast method to control the pollution levels. The mining is carried out by mechanized methods and generated large number of employments, around 5 lakh people by both direct and indirect means. The state is having 400 million tonnes of iron ore resources. The highest production of iron ore from Goa was 40 million tonnes per annum in the past, but for extraction of everyone tonne of ore approximately 4 tonnes of waste rock has to be removed as the mining has gone deep and extent of each lease area is small. The waste rock comprises manganiferous and phyletic clays which are soft and easily erodible. Further, Goa witnesses heavy rainfall up to 3500mm during SW monsoon every year, thus the surface waters are getting polluted. As said above the lease areas are being small and the generation of waste rock is high. So, the disposal of waste rock and its prevention poses a problem. The analysis of waste rock indicates the waste rock is rich of sand grains which can be extractable by simple crushing and screening method. So, a detailed conceptual plan is necessary to sustain the mining and to tackle the environmental issues which are suggested in this paper.

Mathematical relationship between ultimate pit limits generated by discounted and undiscounted block value maximization in open pit mining

Ultimate pit limit is an important aspect of open pit mining. The optimal ultimate outline determines the tonnage of extractable ore, the volume of waste to be removed, the location of the subsidiary facilities, the location of ore and waste stockpiles, the life time of the mine and the estimated net present value (NPV) of the entire mining operation. Traditionally, there are two major approaches to optimizing the ultimate pit limit. One seeks to determine the ultimate pit using undiscounted profit maximization and the other by determining the optimal mining sequence of all blocks and discounting the value of the blocks. The outline with the highest cumulative NPV will be chosen as the final pit limit. For each of these approaches, different algorithms are presented. The aim of this paper is to present an analytical investigation about the mathematical relationship between sets of blocks of ultimate pits generated by each of these approaches in an ore body. This investigation is in fact the mathematical proof of the theorem that a discounted ultimate pit is smaller than or equal to the undiscounted pit. The results show that the discounted pit is always a subset of the undiscounted pit.

About the Activity of "Fergana Rare Metal Mining Society"

It was founded in 1908 in Central Asia, in particular in Fergana Valley, in the deposits "Tuya-Moyin" and "Jil Bulak", rare metals mining - in addition to copper, radioactive chemical elements such as uranium, vanadium, radium, provided information about “Fergana rare metal mining society”. The charter of the “Fergana rare metal mining society” was also analyzed. “Fergana rare metal mining society” was founded by railway engineer Sergei Egorovich Palashkovsky and mining engineer Kh.I. Antonovich. During the “Fergana rare metal mining society’s existence from 1908 to 1918, the “Fergana rare metal mining society” mined 820 tons of ore, out of which 655 tons were shipped to St. Petersburg and uranium and vanadium preparations were processed and exported to Germany. In the 70s and 80s of the XIX century - early XX century, foreign investment was widely attracted, especially in the field of copper mining. The article analyzes the fact that the deposits in Fergana are one of the first uranium deposits in the Russian colony and have their own advanced experience.

Characteristics and Affecting Factors of Sinkhole Development in Cho Don Area, Bac Kan Province, Vietnam

Cho Don district, Bac Kan province, Vietnam is an area where mineral resources are very concentrated. The most potential minerals in the areas are iron, lead, zinc ores and construction materials with large reserves. The mines that have been explored and have large mineral reserves are mines in Bang Lung town’s surrounding area with reserve of about 5,032 thousand tons of ores. Na Tum mine in Cho Don district, Bac Kan province started to operate in April 2007, but since December 2007 up to now a number of large and very large sinkholes have been being developed and most likely shall be developed in future. Na Tum mine dewatering had been considered as a main induced sinkhole formation driving factor. Other necessary and sufficient conditions of sinkhole development and natural factors which accelerate sinkhole formation triggering mechanism have been studied and identified in this paper. The results are essentially important for future comprehensive detailed engineering geological, hydrogeolocical and geotechnical engineering investigation plan which is required for sinkhole geotechnical quantitative analysis, based on which both management and engineering measures to prevent the future sinkhole development, to provide sinkhole warning, to remediation of sinkholes, to carry out the local resident resettlement etc. can be developed. This is an important part of sustainable social-economical development and environmental protection in term of induced geohazards related to mining activities.

Blasting and preconditioning modelling in underground cave mines under high stress conditions

Cave mining is an underground mass mining technique. The largest projects, which are known as 'super caves', produce hundreds of thousands of tons of ore per day, which involves large footprints with considerable column height, and have a life of mine of over 20-40 years. These operations are typically located deep, under high stresses and in competent rock masses, making initiation and propagation of the caving process harder to manage. These challenges must be confronted by optimizing the fragmentation of the orebody to achieve smaller size blocks that will result in consistent caving and improved flow of the ore from the drawpoints. To achieve better performance from the drawpoints, preconditioning is applied to fragment and damage the material required to cave. We present a proposed design for preconditioning in underground mines, considering the challenges that these large-scale mines are already facing, based on a comprehensive analysis of current design parameters, case studies, and sensitivity analyses using numerical models.

Monitoring uranium mine pollution on Native American lands: Insights from tree bark particulate matter on the Spokane Reservation, Washington, USA

The uranium boom in the United States from the 1940's to the 1980's was a period of extensive uranium mining on Native American lands. However, detailed environmental investigations of the resulting uranium pollution are sparse and typically ignore contributions from airborne particulate matter. The Midnite Mine is a 350-acre inactive open pit uranium mine located on the Spokane Indian Reservation in eastern Washington that operated from 1954 to 1981. Approximately 2.4 million tons of ore and 33 million tons of waste rock were left behind in stockpiles and have also been utilized as gravel on access and haul roads. Although the Midnite Mine is now a Superfund Site, and governmental investigations of water and soil contamination have been done, no investigations of airborne particulate matter pollution have been conducted. This study applies tree bark from 31 Pinus ponderosa trees as a biomonitor of this airborne particulate matter. Bulk trace elemental analyses via inductively coupled plasma – mass spectrometry (ICP-MS) of tree bark show that U is the most abundant trace element of interest present up to 232 ppb. Other metals that are of potential human health concern include Th, Pb, and As which are present at 20 ppb, 104 ppb, and 20 ppb respectively. Calculated geoaccumulation indices determine these metals to be at high (U), moderate (Th), and low (Pb and As) levels of contamination. Detailed scanning electron microscopy (SEM) investigations of particulate matter from the surface of tree bark confirm that U and Th-bearing particulate matter exist in the <PM10 size fraction while geospatial analyses indicate that uranium, thorium, and arsenic contamination are centralized along the Midnite Mine access road and at the nearby Dawn Mill where ore was further processed. Combined, these findings indicate that the nature and distribution of historic airborne particulate matter from the Midnite Mine and Dawn Mill provide context for potentially understanding past and current illness on the reservation. In addition, much needed context for future health and environmental studies for both local and national Native American populations is provided.

Geostatistics-block-based characterization of heterogeneous rock mass and its application on ultimate pit limit optimization: a case study

For an open-pit mine, the slope must remain stable throughout the life of mining operation, and it follows that an optimized ultimate pit limit (UPL) should have the slope stability commensurate with economic benefit. In the Shuguang gold and copper mine, a geostatistics-block-based method is used to characterize the heterogeneous mechanical properties of rock mass. Then, the detailed slope stability analyses for four possible slope configuration designs using heterogenous mechanical parameter block model are performed to determine the steepest safe slope angle, and the steepest safe slope angle is next used for the UPL optimization. Compared with the original UPL assuming the rock mass is homogeneous in the same lithology, the slope angle for the optimized UPL has an average 1° to 6° increase, and the optimized UPL can bring 15.84 million tons of ore and reduce 20.83 million tons of waste rock. The result indicated that the application of geostatistics can make practical use of geotechnical information to improve slope stability, and slope configurations, and thereby optimize the UPL and so bring economic benefit.

The late Miocene Öksüt high sulfidation epithermal Au-Cu deposit, Central Anatolia, Turkey: Geology, geochronology, and geochemistry

The Öksüt high sulfidation (HS) epithermal Au (-Cu) deposit (35.32 million tonnes of ore at 1.22 g/t Au) is located in the Central Anatolian Volcanic Province (CAVP) at the south-eastern segment of the Central Anatolian Crystalline Complex (CACC). The deposit is hosted by large hydrothermal breccia bodies and hornblende-rich basaltic andesite porphyry of the Develidag Volcanic Complex (DVC), which were pervasively altered to form silicification and argillic assemblages. The host rocks and mineralization are covered by post-mineral pyroxene-rich basaltic andesite of DVC. The volcanic rocks have a medium-K calc-alkaline affinity similar to regional volcanic rocks of the DVC. The gold and copper mineralization is mostly associated with quartz and primary sulfides hosted by silicification and argillic alteration particularly quartz-alunite cemented breccias and subsequently enriched by supergene oxidation processes causing destabilization of the primary sulfides.The LA-ICP-MS zircon U-Pb geochronology indicates a range of 206Pb/238U ages between 5.4 and 6.0 Ma with an average mean age of autocrystic grains of 5.73 ± 0.06 Ma for the host hornblende-rich basaltic andesite porphyry and 5.67 ± 0.07 Ma for the post-mineralization basaltic andesite cover. The high-precision CA-ID-TIMS analysis of the youngest grains from both rock types defines ages 5.778 ± 0.013 Ma and 5.700 ± 0.019, respectively, which bracket the ore mineralization between 5.791 and 5.681 Ma, including the error uncertainty. Consequently, the Öksüt gold mineralization took place in a very narrow lifetime of 78 ± 32 Ka. This is the youngest age for an economic gold deposit in Turkey.

Non-Gaussian Copula Simulation for Estimation of Recoverable Reserve in an Indian Copper Deposit

The present study developed a geostatistical simulation technique based on non-Gaussian copula for recoverable reserve estimation considering support effect of a well-known open-pit mine of a copper deposit in India. The focus was to examine the efficacy of copula-based simulation model in recoverable reserve estimation. It was assessed by comparing three selectivity curves like grade–ore tonnage, grade–metal tonnage and ore–metal tonnage of reserve constructed using the copula-based simulation, disjunctive kriging and multi-Gaussian kriging with respect to the production data of blasting. The results informed that the copula-based simulation technique provided better accuracy than the nonlinear geostatistical techniques of disjunctive kriging and multi-Gaussian kriging. The root mean square error and standard error analysis indicated that the copula-based simulation model provided more accurate estimates compared with disjunctive kriging and multi-Gaussian kriging. All the three techniques yielded biased estimates, but the least bias was attributed to the copula-based simulation technique.

Methodology for the Estimation and Classification of White Marble Reserves

In the marbel industry, ore reserves refer to the produced tonnage of orthogonal prismatic, defect-free blocks of commercial sizes and of marketable aesthetical quality. On the contrary, in the mining sector, ore reserves refer to the tonnage of ore with grade above some cut-off value. While in the mining sector ore reserves are reported according to existing standard guidelines, the marble industry has no standard method. The purpose of this paper is to present a method to estimate white marble reserves using drill coring or borehole wall logging and limited joint mapping on exposed marble surfaces. Data pertaining to fracture orientation, fracture frequency, size, and whiteness were collected. The primary phase of marble characterization is followed by the discretization of the orebody into a three-dimensional array of orthogonal prismatic blocks, where the marble whiteness and fracture intensity are attribute values. Both these variables are considered to obey continuous and not discrete probability distributions. The discretization is performed using computer aided design and the Block Kriging techniques. An outline to improve the description of a given marble ore deposit at the exploration stage is presented for the prediction of marble block size distribution from joint count measurements along diamond drill cores. An example case of a white dolomitic marble quarry is used to demonstrate these methods.

Vehicle accident risk assessment in mines

The problem of ensuring the safety of modern production is one of the most important. An increase in mineral production and the development of technology lead to an increase in the number of mechanized equipment used in mining enterprises. Transport, in turn, is one of the main factors leading to industrial accidents. The article analyzes the risk of such accidents at large mining enterprises. The share of accidents related to the use of vehicles increases every year and makes up about 25 % of the total number. A mine with annual production of more than 5 million tons of ore was selected as a research model. The risk probability of vehicle accidents in the absence of traffic control was determined. A technique that allows you to take into account the available statistical data on the frequencies and probabilities of possible adverse events, which may be the cause of accidents with vehicles, was chosen to assess risks. Cases of transportation of both ordinary and dangerous goods are considered. The authors determined the probability of accidents for various scenarios and types of vehicles when transporting conventional and dangerous goods, taking into account the lack of traffic light regulation at the interfaces of mining operations. In each case, several possible scenarios for the occurrence of accidents are analyzed. The results of calculations of vehicle accident risks in case of fire and explosion of dangerous cargo are presented. The probability of a collision between one vehicle and another, as well as the probability of a collision when traffic flows on the interfaces, has been established. As a result of the calculation, it was found that the risk of accidents in all cases was acceptable even taking into account deviations from the requirements of the Federal norms and rules in the field of industrial safety. The effectiveness of traffic control to reduce the risk of accidents is planned to be evaluated in further studies. This method allows us to assess the risks of accidents when justifying the safety of mines.

Enrichment of REE and HFSE during the magmatic-hydrothermal evolution of the Baerzhe alkaline granite, NE China: Implications for rare metal mineralization

The Baerzhe Cretaceous (123.7 ± 0.9 Ma) peralkaline pluton underwent extensive fractional crystallization and extreme enrichment in incompatible elements including rare earth elements (REE) and the high field strength elements (HFSE). It is one of the largest resources of rare metals in China, containing approximately 100 million tons of ore with an average grade of 1.84 wt% ZrO2, 1.00 wt% REE2O3 (34% heavy rare-earth oxides), and 0.26 wt% Nb2O5. Zr, REE and Nb are mainly hosted by hydrothermal minerals, such as zircon, hingganite-(Y), monazite-(Ce), polycrase, pyrochlore, fergusonite and columbite. An integrated investigation of field geology, mineral textures and compositions of minerals and host granites was carried out to examine the evolution of the Baerzhe pluton and the roles of magmatic and hydrothermal processes in concentrating REE and HFSE. Most minerals in the intensively altered subsolvus granite show secondary textures or replaced pseudomorphs, such as the replacement of arfvedsonite by aegirine, zircon dissolution and reprecipitation, and the replacement of monazite-(Ce) and polycrase-(Y) by hingganite-(Y). Compositions of the key economic minerals and changes with respect to these alteration stages reveal evidence that hydrothermal alteration played a role in the mineralization of the pluton. In-situ analyses and element mappings suggest that large volume of metals were remobilized and redistributed during hydrothermal replacement, such as the replacing of monazite-(Ce) and polycrase by hingganite-(Y). It is suggested that subsolidus re-equilibration and hydrothermal alteration, in addition to magmatic fractionation, is critical for further concentrating REE and HSFE in peralkaline granitic systems.

Application of geostatistics to complete uranium resources estimation of Rabau Hulu Sector, Kalan, West Kalimantan

Kalan is one of the focus areas for uranium exploration in West Borneo that conducted by BATAN. Situated in the central part of Kalan, previous works in Rabau Hulu Sector consisted of surface geology and radiometric anomaly mapping, trenching, drilling, logging, and conventional uranium resource estimation. Nevertheless, the complete resource estimation of the previous work was still using 2D modeling, and the latest one using 3D modeling is a method-application case study in one orebody. To increase the confidence level and completing the uranium resource estimation of all orebodies in this sector, a geostatistical estimation with 3D orebody modeling using SURPAC mine planning software was conducted in this paper. Gamma-ray log data from 32 drill holes were collected and then interpreted to obtain uranium grade-thickness data. Based on the correlation of grade-thickness data according to surface orebody orientation, the orebody 3D modeling was done. It resulted in 26 orebodies with one control system of lithology as the mineralization only taken place in the quartzite unit. This 3D model then used as a constraint for block model with 4x4x2 m block size and 0.25x0.25x0.125 m minimum block size. Block model calculation was performed using ordinary kriging which generated the kriging efficiency attribute for the determination of the resource category. Within 25 meters searching radius, the calculation resulted in 408, 480 tons of ore, while total uranium resource was 268 tons of uranium with 677 ppm average grade. There were 214 tons of uranium (79%) categorized as measured while the other 54 tons of uranium (21%) categorized as indicated.